kitaly venance d. - blast design optimization to improve material fragmentation-complete report
TRANSCRIPT
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
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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
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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).
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 2
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.
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 3
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.
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 4
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.
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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.
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 6
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
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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.
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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]
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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)
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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
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 11
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
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 12
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).
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 13
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.
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 14
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
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 15
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 ($/ft) = [(Cown + Cop)BL(hr) +BC]/BL(ft)
Where; Cown is cost per hour to own the rig,
Cop is cost per hour to operate the rig
BL is bit life in hours or in feet and
BC is bit cost (Hartman, 1996).
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 16
The cost to own includes taxes, interest, insurance, amortization, and depreciation. The cost
to operate includes labour, fuel, and parts and supplies, such as tires and drill steel. Labour
costs, also applied to blasting costs, include base salary plus benefits. Benefits, which range
from 30 to 40% of base salary, include insurance, health care, pension, and vacation.
Incentive pay, as a percentage of base salary, is often provided when productivity increases
over a predetermined average. Productivity is measured as feet (meters) drilled for the drill
crew or loaded and shot per employee-shift for the blasting crew. Blasting costs comprise
explosives, boosters and primers, initiation systems, and other expendables. Labour costs
include the hours spent by the blasting crew to handle and transport explosives, load holes,
detonate the shot, and take inventory and prepare paperwork. The cost of bulk loading and
storage equipment is also included (Hartman, 1996).
Blasting costs are directly related to powder factor and the cost per pound of the main
explosive charge. Labour costs can represent 5 to 40% of the total blasting costs, while the
cost of expendable blasting accessories such as primers and initiators is generally less than
20% of total costs. A comparison of drilling and blasting costs for various mining methods is
shown in Table 2.3.
Table 2.3: Comparison of Drilling and Blasting Costs for Various Mining
Methods
Extraction Methods Cost ($/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.
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 17
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.
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 18
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).
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 19
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.
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 20
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
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 21
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)
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 22
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).
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 23
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).
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 24
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;
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 25
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)
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 26
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.
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 27
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%
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 28
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).
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 29
Fig. 4.2(e): Burden optimization model
Fig. 4.2(f): Powder factor optimization model
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 30
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.
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 31
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.
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 32
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.
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 33
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)
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 34
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.
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 35
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.
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 36
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Society of Explosives and Engineers Inc. (2010), World of Explosives [online], Available
from: http://www.explosives.org/index.php/component/content/article?id=69 [accessed 27th
January 2013]
Stagg, M.S. and Nutting, M.J. (1987), Influence of Blast Delay Time on Rock
Fragmentation: One-Tenth-Scale Tests, Surface Mine Blasting, Information Circular 9135,
US Bureau of Mines, Washington, DC, pp. 79–95
Strelec, S., M. Gazdek, et al. (2011). “Blasting design for obtaining desired fragmentation.”
Tehniĕki vjesnik 18(1): 79-96
Van Ormer, H.P. (1973), 7 Rules of Thumb for Blasting Hard Rock, Pit and Quarry, Vol.
66, No. 3, Sept., pp. 72–75
Winzer, S.R., Anderson, D.A., and Ritter, A.P. (1981), “Application of Fragmentation
Research to Blast Design: Relationship Between Blast Design for Optimum Fragmentation
and Frequency of Resultant Ground Vibrations,” Proceedings, 22nd US Symposium on Rock
Mechanics, Boston, MA, pp. 237–242
Winzer, S.R., Anderson, S.A., and Ritter, A.P. (1983), “Rock Fragmentation by Explosives,”
Proceedings, 1st International Symposium on Rock Fragmentation by Blasting, Lulea,
Sweden, pp. 225–249
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 39
Yoshikazu H., Kenji M., Yukio K., [n.d.], Detonation Characteristics of Emulsion
Explosives as Functions of Void Size and Volume [online], Shock Wave and Condensed
Matter Research Center, Kumamoto University 2-39-1 Kurokami, Kumamoto 860-8555,
JAPAN. Available from:
http://www.intdetsymp.org/detsymp2002/papersubmit/finalmanuscript/pdf/hirosaki-149.pdf
[Accessed 28th
January 2013]
Ashutosh Mishra (2009), Design of Surface Blasts- a Computational Approach, A Thesis
Submitted in Partial Fulfillment of the Requirements for the Degree of Bachelor of
Technology In Mining Engineering [online], p 12. Available form:
http://ethesis.nitrkl.ac.in/34/2/10505029.PDF [Accessed 28th
January 2013]
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 40
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
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 41
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
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 42
Table 9.1(a) cont….
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
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 43
Table 9.1(a) cont….
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
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 44
Table 9.1(a) cont….
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
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 45
Table 9.1(a) cont….
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
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 46
Table 9.1(a) cont….
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
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 47
Table 9.1(a) cont….
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
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 48
Table 9.1(a) cont….
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
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 49
Table 9.1(a) cont….
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
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 50
Table 9.1(a) cont….
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
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 51
Table 9.1(a) cont….
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
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 52
Table 9.1(a) cont….
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
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 53
Table 9.1(a) cont….
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
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 54
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)
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 55
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)
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 56
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
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 57
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)
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 58
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
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 59
Table 9.2(a) cont….
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
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 60
Table 9.2(a) cont….
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
10-Mar-12 NW-C 5.1 2.5 2.5 CL-RVK SOFT_ROCK 85% 0.77
13-Mar-12 NW-
C&D
3.8 2.5 2.5 CL-RVK SOFT_ROCK 85% 0.71
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 61
Table 9.2(b) cont....
19-Mar-12 SW-C 7.0 2.5 2.5 CL-RVK SOFT_ROCK 80% 0.80
21-Mar-12 NW-C 3.5 2.5 2.5 CL-RVK SOFT_ROCK 88% 0.71
24-Mar-12 NW-C 3.5 2.5 2.5 CL-RVK SOFT_ROCK 85% 0.69
27-Mar-12 NW-C 3.02 2.5 2.5 CL-RVK SOFT_ROCK 85% 0.83
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
06-Apr-12 NW-C 4.2 2.5 2.5 CL-RVK SOFT_ROCK 80% 0.69
11-Apr-12 NW-C 5.0 2.5 2.5 CL-RVK SOFT_ROCK 85% 0.78
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
24-Apr-12 NW-C 3.0 2.5 2.5 CL-RVK SOFT_ROCK 85% 0.70
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
18-Jun-12 NW-C 5.6 2.5 2.5 CL-RVK SOFT_ROCK 85% 0.85
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
28-Jun-12 NW-C 5.2 2.5 2.5 CL-RVK SOFT_ROCK 80% 0.79
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
18-Jul-12 NW-C 7.0 2.5 2.5 CL-RVK SOFT_ROCK 85% 0.83
19-Jul-12 SW-C 10.1 2.5 2.5 BVK HARD-
SOFT_ROCK
70% 0.87
21-Jul-12 NW-C 4.3 2.5 2.5 CL-RVK SOFT_ROCK 85% 0.83
25-Jul-12 NW-C 3.0 2.5 2.5 CL-RVK SOFT_ROCK 85% 0.77
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
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 62
Table 9.2(b) cont....
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
22-Sep-12 NW-B 7.3 2.5 2.5 CL-RVK SOFT ROCK 85% 0.88
25-Sep-12 NW-B 7.7 2.5 2.5 CL-RVK SOFT ROCK 80% 1.13
28-Sep-12 NW-B 7.8 2.5 2.5 CL-RVK SOFT ROCK 80% 0.81
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
12-Oct-12 NW-B 6.8 2.5 2.5 CL-RVK SOFT ROCK 85% 0.98
15-Oct-12 NW-B 6.4 2.5 2.5 CL-RVK SOFT ROCK 85% 0.86
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
22-Oct-12 NW-B 8.2 2.5 2.5 CL-RVK SOFT ROCK 80% 0.75
24-Oct-12 NW-
C,D
4.3 2.5 2.5 CL-RVK SOFT ROCK 80% 0.54
26-Oct-12 NW-B 8.4 2.5 2.5 CL-RVK SOFT ROCK 85% 0.68
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
10-Nov-12 SW-C 7.1 2.5 2.5 NCL-RVK SOFT ROCK 80% 0.82
12-Nov-12 NW-B 6.1 2.5 2.5 CL-RVK SOFT ROCK 80% 0.88
16-Nov-12 NW-B 9.2 2.5 2.5 CL-RVK SOFT ROCK 80% 0.87
19-Nov-12 SW-C 11.4 2.5 2.5 NCL-RVK SOFT ROCK 80% 1.11
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 63
Table 9.2(b) cont....
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
30-Nov-12 SW-C 5.3 2.5 2.5 NCL-RVK SOFT ROCK 80% 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)
© KITALY VENANCE D. (T/UDOM/2009/08472) ~24 January 2014~ Page 64
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)