MASTER'S THESIS Improvement of blast-induced fragmentation and crusher efficiency by means of optimized drilling and blasting in Aitik Ali H. Beyglou Master of Science (120 credits) Civil Engineering Luleå University of Technology Department of Civil, Environmental and Natural Resources Engineering
Improvement of blast-inducedfragmentation and crusher efficiency by
means of optimized drilling and blasting inAitik
Ali H. Beyglou
Master of Science (120 credits)Civil Engineering
Luleå University of TechnologyDepartment of Civil, Environmental and Natural Resources Engineering
Improvement of blast-induced fragmentation and crusher efficiency by means of optimized drilling and blasting in Aitik.
Ali H. Beyglou
Division of mining and geotechnical engineering
Department of civil, environmental and natural resources engineering
Luleå University of Technology
The thesis project presented in this report was conducted in Boliden’s Aitik mine;
thereby I wish to gratefully thank Boliden Mines for their financial and technical support.
I would like to express my very great appreciation to Ulf Nyberg, my supervisor at
Luleå University of Technology, and Evgeny Novikov, my supervisor in Boliden for their
patient guidance, technical support and valuable suggestions on this project. Useful advice
given by Dr. Daniel Johansson is also greatly appreciated; I wish to acknowledge the
constructive recommendations provided by Nikolaos Petropoulos as well.
My special thanks are extended to the staff of Boliden Mines for all their help and
technical support in Aitik. I am particularly grateful for the assistance given by Torbjörn
Krigsman, Nils Johansson and Peter Palo. I would also like to acknowledge the help provided
by Sofia Höglund, Torbjörn Larsson, Jansiri Malmgren and Lisette Larsson during data
collection in Aitik mine.
I would also like to thank Forcit company for their assistance with the collection of the
data, my special thanks goes to Per-Arne Kortelainen for all his contribution.
Finally my deep gratitude goes to my parents for their invaluable support, patience and
encouragement throughout my academic studies.
Luleå, September 2012
Ali H. Beyglou
Rock blasting is one of the most dominating operations in open pit mining efficiency.
As many downstream processes depend on the blast-induced fragmentation, an optimized
blasting strategy can influence the total revenue of a mine to a large extent.
Boliden Aitik mine in northern Sweden is one of the largest copper mines in Europe.
The annual production of the mine is expected to reach 36 million tonnes of ore in 2014; so
continuous efforts are being made to boost the production. Highly automated equipment and
new processing plant, in addition to new crushers, have sufficient capacity to reach the
production goals; the current obstacle in the process of production increase is a bottleneck in
crushers caused by oversize boulders. Boulders require extra efforts for secondary blasting or
hammer breakage and if entered the crushers, they cause downtimes. Therefore a more evenly
distributed fragmentation with less oversize material can be advantageous. Furthermore, a
better fragmentation can cause a reduction in energy costs by demanding less amounts of
In order to achieve a more favorable fragmentation, two alternative blast designs in
addition to a reference design were tested and the results were evaluated and compared to the
current design in Aitik. A comparatively large bench was divided to three sections with three
different drill plans, which led to different specific charges in each section. The sections were
drilled in patterns of 6x9 m, 7x9 m and 7x10 m of burden and spacing; planned specific
charges of the sections were 1.17 kg/m3, 1.02 kg/m3, and 0.91 kg/m3 respectively. Similar to
the current drill plan in Aitik, the section with 7x9 m ( 1.02 kg/m3 specific charge) was used
as the reference for results comparison. The drilling and charging processes were monitored
carefully and the post-blast parameters were measured accordingly. Laser scanning was used
to measure the swelling of the sections and two different methods of image analysis were
utilized to evaluate the fragmentation of the rock for each section. Drilling log data (MWD)
were analyzed to evaluate the hardness of the rock; energy consumption log of the crusher
was also analyzed and all the data was collected in a single database. VBA (Visual Basic for
Applications) programming language was embedded within data spreadsheets to correlate the
mentioned data to the coordinates of the rock by means of Minestar logs, which include both
timestamps and coordinates of all machinery e.g. shovels and trucks.
The results of the test show significant improvements in fragmentation and oversize
material percentage in the section with 6x9 m drill plan (1.17 kg/m3). The advantage of 6x9
m plan was confirmed by 52% higher swelling, 66% lower oversize material and 26% lower
crushing energy compared to the reference section. The section with 7x10 m drill plan (0.91
kg/m3) also showed theoretically acceptable results; however, the deviations from reference
were not as large as formerly mentioned section. The swelling had a decrease of 8%
compared to the reference section and the percentage of oversize material and crushing
energy were increased by 16% and 2% respectively.
Presented results are based only on technical aspects and do not include the costs of
drilling and charging. Thus, in order to evaluate the drill plans in practice an economical
evaluation of the sections should be conducted. Also a confirmation test with more accurate
geology explorations is recommended.
Finally, upon the request of Boliden Mines, a short report on the usage of Air-decking
technique in Aitik is enclosed as an appendix. The report includes a brief introduction to air-
decking and discusses practical solutions to apply this technique in Aitik.
TABLE OF CONTENTS 1 INTRODUCTION ............................................................................................................... 1
1.1 Aims and objectives .................................................................................................. 2 2 THEORY AND BACKGROUND ...................................................................................... 4
2.1 Rock breakage by blasting ........................................................................................ 4 2.2 Mechanical properties of rock mass .......................................................................... 6 2.3 Bench blasting ........................................................................................................... 8 2.4 Particle size distribution .......................................................................................... 10 2.5 Influence of blasting on downstream processes ...................................................... 12
3 AITIK MINE ..................................................................................................................... 13 3.1 Drilling and blasting ................................................................................................ 14 3.2 Loading and hauling ................................................................................................ 16 3.3 Crushing and grinding ............................................................................................. 17
4 TEST BLAST .................................................................................................................... 18 4.1 Description .............................................................................................................. 18 4.2 Data quality and constraints .................................................................................... 22
5.5 Crusher efficiency ................................................................................................... 46 6 DISCUSSION AND CONCLUSIONS ............................................................................. 48 7 RECOMMENDATIONS AND FURTHER STUDIES .................................................... 51 REFERENCES ........................................................................................................................ 53 ADDITIONAL BIBLIOGRAPHY ......................................................................................... 54 APPENDIX I: Fragmentation analysis of FragMetrics software APPENDIX II: An introduction to Air-decking in Aitik
Open pit mining is one of the most utilized methods of ore extraction worldwide. Cost-
effectiveness, mechanical ease and safer environment are some of the advantages of open pit
mining over other mining methods, in addition to that, its potential for large production
volumes and low cost of recovery allows low-grade ore bodies to be extracted feasibly.
Drilling and blasting is, by far, one of the main operations in open pit mines, affecting
the total revenue of the mine to a large extent. Pre-blast costs, such as drilling and explosive
expenses, are directly influenced by blast design; post-blast parameters are also affected by
the outcome of the blast; secondary blasting, loading and hauling, crusher throughput, and
grinder efficiency are related to blast-induced fragmentation of the ore (Nielsen and Lownds
1997, McKee et al. 1995).
Blasting is the most energy efficient stage in the comminution process. According to
Eloranta (1997), blasting has an energetic efficiency of 20% to 35%, which is relatively high
compared to respectively 15% and 2% efficiency of crushing and grinding. High efficiency
offers the blasting stage a strong potential for optimization of the overall comminution
process, however, the operations included in the comminution process were treated
individually for a long time and the optimizations were merely limited to the outskirts of each
operation. A more recent approach to optimization is called ‘Mine-to-Mill’, provided by
Julius Kruttschnitt Mineral Research Centre in 1998 (JKMRC 2012). Mine-to-Mill, in short,
is ‘an approach that identifies the leverage that blast results have on different downstream
processes and then optimizes the blast design to achieve the results that maximize the overall
profitability rather than individual operations’ (Grundstrom et al. 2001). According to Mine-
to-Mill concept, blasting should be designed in a way that satisfies the overall requirements of
the comminution process, including haulage, crushing and grinding altogether.
Research works by Eloranta (1997), Kojovic (2005), Ouchterlony (2003 and 2005)
and Ouchterlony et al. (2010) show a meaningful relation between blast properties and
efficiency of crushing and grinding. Therefore the optimization of blasting should not only
include the size distribution of blasted rock, but also consider the crusher throughput and
grinder energy consumption.
The importance of blasting has also urged the necessity of reliable monitoring systems
to develop. Fragmentation is a key factor in the comminution process and image analysis has
been the most utilizable method of fragmentation measurement so far (Chiappetta 1998).
Developments in that field have led to systems able to measure the fragmentation
continuously during mining; such systems, together with well-calibrated measurements of
throughput and energy consumption in crusher and grinder, provide an appropriate database
to optimize the process.
1.1 Aims and objectives
The main goal of the current project is to find an economically viable alternative blast design
to provide an improved fragmentation as well as an increase in the energy efficiency of the
Constant efforts are being made in the Aitik mine to optimize the production process;
accordingly, blast-induced fragmentation is of significant importance as a major role player in
such optimization. Boliden Mines implemented an expansion of operations project (Aitik 36)
during the period of 2006 to 2010. Pushbacks at the southeastern, northeastern and western
sides of the pit resulted in trebling of ore reserves from 200 Mt to 600 Mt, as well as an
extension in mine life from 2016 to 2025 and the ability to excavate down to 600 m depth. In
2010 a new modern processing plant has been inaugurated in accordance to Aitik 36
expansion project, aiming to increase the annual production up to 36 Mt until 2014; but
presently, the crusher, which is directly influenced by fragmentation of blasted rock, is a
bottleneck in comminution process.
Large number of oversized boulders requires high cost and effort for secondary
blasting; in addition to that, accidental throw of oversized boulders in the crusher opening
causes downtime in the crusher, which creates a bottleneck in the production. A solution to
such problem is to modify the blasting parameters in a way that improves the size distribution
of the fragmented rock. The alteration in parameters should not only result in fewer boulders
but also in an increase in the energy efficiency of the crusher.
2 THEORY AND BACKGROUND
2.1 Rock breakage by blasting
The entire blasting act takes only a few seconds in scale of time. However, several events take
place in different segments of those seconds. Once initiated, the explosive 1 releases an
enormous amount of energy through chemical reactions, resulting in high-pressure gases in
the blast hole which can amount to and exceed 10 GPa. The high pressure of gases is not, in
and of itself, the only cause of the breakage; the rapidity of the reaction plays the leading role
(Langefors and Kihlström 1967).
Upon initiation, the reaction advances at a rate (Velocity of Detonation, VOD) of
approximately 2000-6000 m/s throughout the explosive. Considering the 15-20 m length of a
normal blast hole, one can easily realize that the reaction takes place within thousandths of a
second. The rapid reaction leads to an almost instantaneous pressure rise in the hole, which
produces a shockwave in the rock, traveling at a speed of 3000-5000 m/s. The high pressure
expands the walls of the hole and the area adjacent to the drill hole shatters as a result of vast
amounts of tangential strains and stresses. The shattered area around the hole, with rose
shaped cracks towards outside of the hole, is the first platform for fracturing (Esen et al. 2003).
See figure 2.1.
1 In this report, the word “Explosive” refers to non-military, civil explosive materials used in mining industry.
Figure 2.1: Fractured area around the blast hole, the so-called Rose of cracks, After Esen et al. (2003).
The shockwave travels at such high speed that the initial cracks form within a few
milliseconds. According to wave propagation concept (Hustrulid 1999), the positive pressure
of shockwave falls rapidly to negative values, which implies a change from compression to
tension. Since rock is generally more resistant to compression than to tensile strain, the initial
radial cracks are the results of tensile forces acting on the area around the hole.
During the first stage there is practically no breakage in the rock other than the radial
cracks. The main breakage occurs after the shockwave reaches the free face of the rock and
reflects as a tensile wave, such phenomenon gives a rise to the tensile strains and
consequently extends the cracks throughout the rock. This stage is called Scabbing (Langefors
and Kihlström 1967).
The scabbing and radial cracks are both effects of the shockwave; the last stage of
breakage is under the influence of pressurized gases produced by the blast; this stage is
considerably slower than the first two. The high-pressure gases in the blast hole, kept inside
by the stemming, pressurize the borehole and apply a radial compressive stress perpendicular
to the borehole; the compressive stress is large enough to initiate new cracks and extend the
existing cracks. The crack expansion outspreads through the rock and results in breakage. The
overall displacement of the rock mass prior to gas action is very little; the gases not only
extend the cracks, but also exert a pushing force to move the broken rock forward (Langefors
and Kihlström 1967).
A successful, complete breakage takes place when the amount of explosive and the
geometry of the blast e.g. burden, spacing, height, are balanced in a way that the cracks
expand all the way to the free face and gases push the rock forward to form a well-swollen
pile; so it is critical to find the appropriate proportions of these factors based on the rock
strength, fracturing, and explosives characteristics in order to reach an adequately broken and
swollen rock pile.
2.2 Mechanical properties of rock mass
The strength of rock can be defined by many parameters, e.g. compressive and tensile
strength. The large difference between intact rock strength and rock mass strength cause many
uncertainties in large-scale mining activities. Figure 2.2 shows the difference between rock
mass and intact rock; the mechanical behavior of rock mass is heavily affected by
discontinuities, varying in a wide range from microscopic cracks to regional faults. In
addition to that the direction of the load and confinement conditions effect the behavior of the
rock mass; high confinement pressure turns brittle failure to ductile and due to closure of
micro cracks Young’s modulus increases (Brady and Brown 1993).
Figure 2. 2: Schematic difference of intact rock and rock mass, after Scott et al. (1996).
In addition to that, the failure of rock includes an element of creep, which means the
loading rate effects the strength of the material (Bergman 2005). Crushing and grinding of
rock are performed through static loading of the rock. However, blasting exposes the rock to
both static and dynamic loading due to rapid explosive reactions. Such dynamic loading
exposes the rock to high loading rates, which results in higher compressive strength (Persson
et al. 1994) and increased Young’s modulus of the rock (Bergman 2005).
The complexity of rock mass characterization for blasting purposes leads to the
conclusion that analytical solutions are not possible. As of yet, empirical measures have been
the most useful tools to classify rock masses. The rock constant, c, is one of the most utilized
tools in blast designs. Rock constant is a measure of the amount of explosives needed to break
one cubic meter of rock and it is determined by controlled trial blasts in a vertical bench
(Langefors and Kihlström 1967).
2.3 Bench blasting
Achieving a well-distributed particle size is the main goal of blasting, so that the rock can be
handled efficiently in post-blast processes, e.g. loading and crushing. The outcome of blast is
influenced by several parameters; mechanical properties of rock mass, geometry of blast holes,
type and amount of explosives, initiation pattern and delay times are some of the key factors
in blast design. A brief terminology of bench blasting geometry is presented in Figure 2.3.
Figure 2.3: Bench blast geometry and terminology, Bergman (2005).
Specific charge, in addition to the geometry, is a key factor in bench design
(Langefors and Kihlström 1967). The specific charge, q, represents the explosives
consumption per cubic meter of rock (or per tonne rock). Specific charge varies based on
explosive and rock mass characteristics, see equation 2.1.
q: Specific charge (kg/m3)
Q: Total explosive per hole (kg)
B: burden (m)
S: Spacing (m)
H: Bench height (m)
Burden, B, is the distance between the rows and spacing, S, is the distance between
the holes in a row. Several empirical equations are provided in the textbooks for calculation of
the burden; Langefors and Kihlström (1967) provided a well-known formula for calculation
of maximum burden, equation 2.2.
D = blast hole diameter (mm),
p = explosive density (kg/dm3)
E = weight strength of explosive (%)
B = burden (m),
c = rock constant (kg/m3)
f = degree of confinement, 1 for vertical holes.
Other parameters are basically calculated by empirical rules of thumb, such as
(Persson et al. 1994):
3.0max ×= BU (2.4)
)(05.1 UHL += (2.5)
These empirical relationships are usually modified depending on the size and characteristics
of the blast.
2.4 Particle size distribution2
The results of a production blast are mainly presented by fragmentation of the broken rock.
The fragmentation is described in terms of geometrical characteristics of the particles i.e. size,
angularity or roundness. The cumulative size distribution function, CDF, provides a complete
description of the former. It is either obtained from physical sieving of the material, which is
very costly in large-scale blasts, or by non-physical sieving methods such as image analysis.
The CDF is the ‘fraction of mass P passing a screen with a given mesh size x.’(Ouchterlony
2003). Percentage of passing material from each mesh, P(x), varies between 0-100%, see
2 Mainly based on the SWEBREC report by Ouchterlony (2003).
Figure 2.4: Cumulative size distribution curve, Ouchterlony (2003).
Depending on the purpose of the analysis, several distinctive quantities are extracted
from the curve, here follows some:
X50 = a measure of the average fragmentation, i.e. mesh size through which half of the
material passes, X50 is a central production measure.
XN = other percentage related block size numbers in use. N=20, 30, 80, 90 etc.
PO = percentage of fragments larger than a typical size XO. PO is related to e.g. the
handling of big blocks by trucks or the size of blocks that the primary crusher cannot
PF = percentage of fine material smaller than a typical size XF.
In large-scale production sites the focus is on the most important of these, which in
Aitik case is PO, due to problems caused by boulders at the crusher feed.
2.5 Influence of blasting on downstream processes
The effect of blasting on subsequent operations has drawn a great deal of attention in recent
years. In the past, the only criterion for blast results was the ability of excavation and hauling
equipment to handle the blasted rock, but mining economy demands high production
capacities as well as efficiency of costly operations. Since crushing and grinding consume
enormous amounts of energy, the effect of blasting on efficiency of these operations is
The effect of blasting on fragmentation is assessed in two different aspects: Seen and
Unseen. The size distribution of blasted fragments is the “seen” part of blasting results, which
can be measured quantitatively by sieving or image analysis techniques. The “unseen” effect
of blasting is the fracture generation within the fragments, these fracture can be classified as
either macrofractures or microfractures. Macrofractures are comparatively large and can be
seen on the surface of fragments; but microfractures are only seen through a microscope
(Workman and Eloranta 2003).
The production and downtime of the crusher are under direct influence of the “seen”
effect of blasting; oversize fragments cause a reduction in primary crusher throughput and
lead to more downtime for clearing the crusher bridging (Workman and Eloranta 2003). On
the other hand, the “unseen” aspects of fragmentation influence the energy consumption of
the crusher. Therefore, it is very important to assess the effect of blasting on the energy
consumption of the primary crusher in addition to the fragments’ size distribution.
The degree of dependency of crushing efficiency on macro and microfractures is not
presently clear; but studies and field tests by Eloranta (1995), Workman and Eloranta (2003),
Ouchterlony (2003) and Ouchterlony et al. (2010) confirms that fracturing of the fragments,
caused by heavier blasting, leads to lower energy consumptions in crushing and grinding
3 AITIK MINE
Aitik open pit mine is situated outside the city of Gällivare in northern Sweden. The orebody
consists of low grades of copper, gold and silver. The production started with two million
tonnes of ore in 1968 and gradually increased to 31.5 Mt in 2011; the production level is
expected to reach 36 Mt in 2014.
The pit is 3 km long, 1.1 km wide and 425 meters deep; the orebody dips 45° towards
west and mainly consists of metamorphosed plutonic, volcanic, and sedimentary rocks with
Investigations by both Sjöberg (1996) and Bergman (2005) on rock strengths in Aitik
show fairly similar results. Muscovite schist is the weakest rock with an average strength of
64 MPa and Pegmatite is the strongest rock with a 141 MPa compressive strength. Biotite
schist and Biotite gneiss were similarly approximated to have strengths of 88 and 121 MP
The utilized method of excavation in Aitik is pallet mining, in which the ore is
removed in form of horizontal slices. The comminution process is shown in figure 3.1; after
drilling and blasting the ore is loaded into trucks and hauled to the in-pit crusher. Once
crushed, a conveyor belt transports the ore to two ore piles that feed the grinding mills. Later
the grinded ore goes through chemical processes and finally the produced concentrate is
transported to smelter by railway.
Figure 3. 1: Processes involved in mining in Aitik, after Bergman (2005).
3.1 Drilling and blasting
The typical drill plan presently used in Aitik consists of 311mm production holes and 127 and
152mm holes for contour blasting. As seen in figure 3.2, production holes are drilled in
accordance to 7m burden and 9m spacing. Contour holes are drilled with 4 and 5 meters
spacing for first and second row respectively, the rows are distant 4.5m from each other and
6m from first production row. Standard benches are 15m high and blast holes are
approximately 2m sub-drilled.
Figure 3.2: Current design of bench drilling in Aitik.
Four Atlas Copco pit viper PV 351 drill rigs are used for production drilling. The rigs,
equipped with GPS and Terrain for Drilling3 system, are of the most advanced blast hole
drills on the mining market. The coordinates of holes are uploaded to the rig and the rig
navigates to the precise coordinates of each hole using GPS. With 56700 kg of bit load and up
to 107.6 m/min of air at 758kPa, the vipers provide a high capacity of fast drilling. The MWD
system logs all the drilling data, such as torque, penetration rate, feed pressure etc. The logged
data will be analyzed and used for interpretation of the properties of the penetrated rock mass
e.g. hardness, fracturing and hydraulic conditions.
Emulsion explosive is used as the main charge in blast holes; it has an average VOD
of 5700 m/s and is of 1350 kg/m3 density. The emulsion matrix is made in a nearby factory.
Special trucks carry the emulsion matrix, Ammonium Nitrate, diesel and water to the benches.
3 Formerly known as AQUILA™
The trucks are equipped with a system to mix the matrix with diesel and AN beforehand
charging. The temperature of the contents, as well as the mixture proportions and volume of
the explosive filled in each hole are set through the computerized system of the mixing trucks.
The current specific charge of production blasting in Aitik is 1.02 kg/m3, which varies from
time to time depending on geology and production requirements.
Nonel Unidet system is used to detonate the holes. Two boosters combined with two
detonators are placed at the bottom of each hole to assure the detonation of the emulsion, the
boosters and detonators are of types Dyno 1.7 and Nonel U-1000 respectively. The coupling
takes place after plugging the holes with about 5.5 meters of crushed stemming material; the
holes are detonated with 176 ms of delay between the rows and 42 ms delay between the
holes in a row.
3.2 Loading and hauling
Four shovels and 30 trucks of various capacities are used to load and haul the blasted rock.
All vehicles are equipped with Minestar system. Minestar is an integrated operations and
mobile equipment management system; Tracking ore and waste, locating the vehicles and
managing the schedule and assignments of the fleet operations are some of the capabilities of
Presently one of the shovels, of type P&H 4100C, is equipped with a camera installed
on the boom. The camera is part of Fragmetrics™ fragmentation measurement system; it
captures photos of the bucket every two minutes. The photos are analyzed with Fragmetrics
image analysis software to estimate the fragmentation curve of the loaded rock.
3.3 Crushing and grinding
The in-pit crusher, Allis-Chalmers Superior 60-109, does the main part of primary crushing of
the main pit rock. It is situated at the 165 m level and consists of two primary gyratory
crushing stations as well as overland conveyors and feeders. The system has a capacity of
8000 t/h and will transport the ore 7 km. At downtimes, or during maintenance periods, two
crushers on the surface are used. The main crusher’s opening is 152 cm in diameter and the
lower part of the mantle has a diameter of 277 cm. Depending on ore properties, the coarsest
boulders’ size after crushing varies in range of 35 to 40 cm.
The crushed ore is transported to two stockpiles on a conveyor belt. The total capacity
of the stockpiles provides 16 to 20 hours of full production in the mill, i.e. 50000 tonnes.
The grinding process is operated through five grinding lines in three grinding sections.
Each grinding line consists of a primary autogenous mill and a secondary pebble mill. The
process is a close chain, a screw classifier feeds the coarse material back to the primary mill
and pebbles are extracted from the primary mill and fed to secondary mill, Figure 3.3.
Figure 3. 3: Milling process of the ore, Bergman (2005).
Finally the grinded ore is transformed into concentrate by processes of flotation,
thickening, dewatering and drying. The concentrate is then transported by railway to
Rönnskär smelter in the city of Skelleftehamn.
4 TEST BLAST
A production bench, named S1_210_13, with a volume of 774000 m3 was assigned for the
test; it was situated at the western wall of the pit at 210 m level. The bench was divided to
three smaller sections, of which one was used as reference. Hereinafter the letters A, B and C
are referred to these sections. Figure 4.1 shows the bench; a drilling plan of 6 m burden and 9
m spacing was assigned to section A. Section B, with the currently used drill plan (7 m
burden and 9 m spacing), was the reference for further comparisons; the middle section was
chosen as reference in order to minimize the effect of geological discontinuities in the blast
results. Section C was ascribed a wider drill plan with respectively 7 and 10 m of burden and
The test bench consisted of a total number of 668 holes; average planned depth of
holes was 16.2 meters of which 1.2 m was sub-drilling. The planned specific charges were
1.17, 1.02 and 0.91 kg/m3 for sections A, B and C respectively. A stemming length of 5.5
meters was also planned for all holes; gravel of size 5-8 cm was used as stemming material.
During the drilling, some practical issues regarding neighbor benches and machinery
led to a change of plans; a part of section A was omitted from the test bench and scheduled
for the next round. The omitted part is shown with red rectangle in Figure 4.1.
Figure 4. 1: Test bench and drilling patterns.
The bench mostly consists of Muscovite and Biotite Gneiss. Two dykes of Biotite cut
through the bench diagonally, a large part of an Amphibolite Gneiss dyke also cuts the
southeastern edge of the bench (Figure 4.2). Although the geology map does not show any
Pegmatite dyke within the bench area, their existence cannot be discarded for sure as the
precision of explorations are not so high.
Figure 4. 2: Geology of the test bench.
Except the drill pattern, all the blast parameters such as emulsion density, hole depth,
sub-drill etc. were kept unchanged in order to have a parametrically controlled test. Drilling
and charging processes were also controlled and monitored to avoid fortuitous errors.
Nonel Unidet system was used for initiation of the blast. The initiation plan is shown
in Figure 4.3. The southward direction of the blast was decided based on the direction of rock
structure as well as loading availability. The blast initiated at the northeast corner of the bench
spreading towards southwest. A delay of 176 ms was used between the holes and each row
was delayed 42 ms from previous one. As a result of smaller free face for the blast at the
northwestern part of the bench, shorter delays of 67 and 109 ms were introduced for a
smoothly swollen rock pile. To prevent detonation failures, two detonators and two boosters
were used for each hole.
Figure 4. 3: Initiation pattern of the blast holes; the initiation starts at the upper left part in the figure.
The blast itself was filmed using a high-speed camera so the initiation of all holes
could be confirmed. Once blasted, the surface of the bench was laser-scanned to evaluate the
swelling of the rock in three sections.
The rock was photographed continuously during loading. The Fragmetrics camera,
installed on the boom of one of the shovels, photographed the bucket every two minutes. A
10-megapixel camera, Nikon D3000 equipped with a 300 mm tele lens, was also used to
manually photograph the bed of loaded trucks.
To estimate the fragmentation of blasted rock two different softwares were used:
FragMetrics™ and Split-Desktop™. Fragmetrics was used to analyze the photos taken by its
associated camera; manually taken photos were used for Split-Desktop. The Fragmetrics
camera was calibrated in accordance to the width of the shovel bucket (460 cm); no further
adjustments were needed since the position of the camera was fixed. However, manually
taken photos could not be shot from exact positions, so photos were taken from a suitable
level depending on the location of each truck, pictures are shot from an angle which
minimizes the optical distortions in the rock pile. The width of the flatbed of trucks was used
as the scale for size distribution analysis.
To obtain a consistent correlation between the crusher throughput and fragmentation,
the measurements required to be obtained from the identical rock. In order to track the ore
from the bench to the crusher, timestamps were attached to photos. The manual camera, as
well as Fragmetrics camera, was synched to the clock of Minestar system, so the photos could
be linked to certain times of loading and crushing.
By estimating the average time delays for the ore between unloading and entering the
crusher, the crusher throughput could be linked to a certain set of photos. Thus, a size
distribution curve is available for each value of throughput. The location of shovel, obtained
from Minestar, shows the location of the loaded ore and correlates this relationship to one of
the three sections of the bench.
To avoid mixtures of the ore in the crusher, all the trucks were assigned to load the ore
only from the test bench during data collection, so the crusher was only fed with the ore from
the test bench and no ore mixing took place.
4.2 Data quality and constraints
A variety of data sets are involved in the process of correlating pre-blast measurements to
post-blast parameters, each data set is obtained from a different source, including its
systematic errors. In addition to that, each source has a specific chance of failure, which
causes loss in the data and/or large errors. Thus, the necessity of an appropriate analysis
strategy based on the availability and quality of the data is inevitable. Following sections
briefly describe some of the sources, their systematic errors and availability of their data.
4.2.1 Drilling (MWD)
Drill rigs in Aitik are equipped with Aquila DM-5 drilling management system for precision
drilling through GPS positioning. MWD (Measure While Drilling) is a part of this system that
collects and archives the drilling parameters while drilling each hole. Although this data is not
fully used in production yet, it has been studied and evaluated several times and there is no
doubt in its usefulness.
Several parameters are included in MWD logs; some are independent parameters,
others depend on the geological and geotechnical properties of the rock mass. Depth, time,
rotation speed and feed force are independent parameters while penetration rate, torque,
vibration and air pressure are parameters that depend on rock mass characteristics. All these
parameters implement a systematic error in the measurements, but since there are no reference
measurements to evaluate the errors, one cannot quantify these errors in a numerical manner.
One of the most critical parameters in drill measurements is depth, due to the fact that
all other parameters are recorded along the depth of the hole. The depth of the hole also
decides start and end of drilling. In this project the depths of 215 holes right before and 1 hour
after charging were measured manually to control the MWD measurements of depth, the
results are presented in 4.2.2 and show an acceptably low error. However, a calibration of the
depth measuring system will improve the accuracy to a large extent.
Another issue with MWD data is its multi-dimensional nature. Each parameter is
recorded along the depth of the hole; in order to correlate a parameter to XY coordinates for
several holes, one should eliminate one of the dimensions. In other words, only one value can
be assigned to the hole in order to analyze the parameters horizontally. Usually penetration
rate (PR) is the governing parameter for analysis, which is also dependent on other drilling
parameters such as feed force. A solution to that is using a calculated index that includes
several variables. Specific energy (SE) is a concept that represents the work done per unit
excavated. The concept was introduced by Teale (1964) and has been evaluated by
Schunesson (2007). Teale (1964) introduced the following equation for specific energy:
SE = Specific Energy [N.cm/cm3]
F = Feed Force [N]
A = the cross-section area of the drill hole [cm2]
N = Rotation Speed [RPM]
T = Torque [Nm]
P = Penetration Rate [m/minute]
In this report both penetration rate and specific energy are presented. By introducing
specific energy alongside penetration rate, the errors in PR are eliminated or minimized; but
SE is still a parameter measured along the depth. In this study an average value within a fixed
depth of the hole has been considered as a median value for horizontal analyses. Figure 4.4
shows a sample MWD data for one of the holes. Custom VBA codes were written to analyze
the data; the average values have been calculated for the depth of the hole in between the two
red lines (the lines are unique for each hole). In this way the fractured part near the surface
(Sylta) as well as the low values at the bottom of the hole are ignored and errors due to
median calculation are eliminated and the median value can be considered as a suitable
representative of rock quality.
Figure 4. 4: MWD analysis of a sample hole (Hole #264).
The final issue with MWD data, for which no solution exists, is loss of data due to
mine network failure. As it will be shown in following chapters, MWD measurements were
not available for some parts of the bench. Fortunately the loss was not extreme, so there was
still enough data available to evaluate the test; but obviously more data is more favorable as it
helps to acquire more accurate results.
Although highly mechanized trucks are utilized for charging process in Aitik, errors still
occur due to rock fractures and human error. Manual measurements of the depth of 215 holes
before and after charging show high deviations resulted by operators’ error (Figure 4.5). The
measurements were conducted before and 1 hour after charging, the 1-hour delay was
sufficient for the expansion of emulsion explosive and it was short enough to avoid the
leakage of emulsion into joints and fractures.
Each point in Figure 4.5 represents the deviation of drilling and charging depths from
the planned depths of one blast hole, the closer the point to the 0 circle, the smaller the error.
Since the measurements were focused on a specific area on the bench, any comparison
between the errors of three sections would not be judicious.
Figure 4. 5: Drilling and charging errors of 215 blast holes.
Despite the problems caused by groundwater during and after drilling, figure 4.5
shows acceptably low errors in drilling depths. However, charging errors are distinctively
larger; a few holes were completely filled with emulsion, these overcharged holes were
blasted without stemming and caused fly-rock and air-blast risks. Some holes were similarly
undercharged; the undercharged holes result in low specific charge in some areas, which lead
to coarser fragmentation and more oversize boulders.
Shovel logs are of great importance in evaluation of fragmentation and crusher performance.
The logging system records all the parameters every few seconds. The GPS positioning
system installed on the shovel records three different coordinates at each recording, one for
the shovel itself, one for the shovel bucket and the last for the digged area. The important
coordinates are digging and bucket coordinates; further analyses regarding fragmentation and
crusher throughput are correlated to the rock coordinates using digging coordinates. The
bucket position log is used to evaluate the digability of the fragmented rock; the bucket
coordinates are linked to the time taken to load the bucket, which is the standard measure of
Depending on the GPS system and satellite visibility, there is an error of about 1 to 15
meters associated with GPS measurements, especially in higher latitudes (Aitik mine is
located at approximately 70 degrees latitude). By graphing the position of the shovel and
comparing it to the laser-scanned map of the bench it was revealed that the accuracy of the
GPS was more than expected. No overlap was observed near the walls and on the edge of the
bench, which shows a reasonably good accuracy.
4.2.4 Trucks (Minestar)
Aitik mine uses Cat® Minestar™ system for material tracking and real-time fleet management.
Logging the trucks activities is one of the capabilities of this system that is used in the current
study. Each truck is identified by a unique ID number, which is used to correlate the
fragmentation (extracted from manually taken photos) to the crusher parameters.
Unfortunately the system was functioning faulty and the truck IDs did not match the
ID signs on the trucks. Providentially the logs included correct data about times and locations
of trucks to extract a representative for each section of the bench. As a solution to faulty
logging of truck IDs, the truck cycle IDs from shovel are used; truck cycle ID is another
unique ID which identifies each cycle of the trucks within the mine. The procedure is
described in detail in 4.3.
An effective method to assess fragmentation is to acquire digital images of rock fragments
and to process these images using digital image processing techniques. In the case of post-
blast fragmentation, this is the only practical method to estimate fragmentation, since
screening is impractical on a large scale.
The Split-Desktop software was originally developed at the University of Arizona; in
1997 the technology was commercially available through a newly formed company, Split
Engineering. The Split software allows post-blast fragmentation to be determined on a regular
basis throughout a mine, by capturing images of fragmented rock in muckpiles, on haul trucks,
or from primary crusher feed. The resulting size distribution data can then be used to
accurately assess the fragmentation associated with different parts of a shot (Kemeny et al.
The basic steps involved in Split-Desktop analysis are acquiring images, pre-
processing the images to correct lighting, defining scales, delineating the images using Split
algorithms and finally correcting the delineation manually and defining fines (Figure 4.6).
The complete manual correction of delineations takes about 30 to 90 minutes per image,
depending on the quality of the image, lighting conditions and presence of dust, fog or other
obstacles and the level of precision in delineation. The software then applies statistical
algorithms to the 2D particle distribution to determine 3D particle volumes. To achieve an
average distribution multiple images should be processed, preferably with different scales.
Figure 4. 6: Delineation of a sample image with Split-Desktop.
Ouchterlony (2003), Kemeny et al. (2002) and Sanchidrian et al. (2006) have
extensive studies regarding the procedures and errors in measurements of fragmentation by
image analysis. Based on their studies and in accordance to the conditions, the sources of
errors associated with the current study can be concluded as follows:
- Sampling error
- Optical distortions
- Manual corrections of delineation
In order to minimize the sampling errors, efforts should be made to acquire evenly
distributed images among the bench area. The entire bench had been photographed during
loading, but unfortunately the limited data from Minestar did not permit an evenly distributed
analysis over the entire bench; as a solution to that three areas on the bench were selected as
representatives for three sections and images were sampled evenly within those areas.
The optical distortions were also mostly overcome by using a tele lens. The long focal
distance of the lens (300mm) minimized the image distortions to a great extent. The photos
were also taken from an approximately fixed angle and two separate scales were defined in
each image, so the scales and perspective of all images are roughly the same.
Split-Desktop results are highly user-dependent, in other words the fragmentation
obtained from an identical image is not the same for two different users. In that regard, only
one user has analyzed all images; the error of different delineating styles were also minimized
by making an efforts forth fairly identical delineation styles through all images. Although
much effort has been made in order to eliminate the errors in fragmentation analysis, the
existence of systematic errors in such process in undeniable.
FragMetrics™ is a fragmentation measurement package provided by Motion Metrics
International Corporation. The package includes a camera installed on the boom of a shovel,
logging and storage devices, and a tablet PC with FragMetrics software to process the stored
The principle of FragMetrics is the same as Split-Desktop. However, the inert position
of the camera against the bucket eliminates the optical errors. In addition to that the
automated essence of the system provides a very useful tool for continuous monitoring of
FragMetrics is a newly developed system. It started operating from January 2012 in
Aitik, so very few experience regarding its results is available. Preliminary evaluations of the
results reveal that in contrary to its advantages to Split-Desktop, the very low resolution of the
camera leads to unreliable results. Figure 4.7 shows a sample image from FragMetrics camera,
the low resolution of the image limits the particles’ visibility down to boulders only.
Figure 4. 7: A sample image and delineation from Fragmetrics system.
A normal image analyzed in Split-Desktop has a resolution of around 2500x1500
pixels, showing a wide range of particles; but FragMetrics images are limited to 380x150
pixels, which is very low comparatively. Such resolution leads to faulty size distribution
analyses. Therefore the FragMetrics size distribution results were not used in evaluations and
discussions of this report; but as the boulders were clearly visible in the images, they
represented a very good source of data for oversize material. The images were used only to
determine the percentage of oversize material in each section of the bench.
As mentioned before, the loading continues round the clock in Aitik and the loaded rock is
from different sources. In order to avoid rock assortment one of the crushers (KR 165) had
been only fed with the rock from the test bench. The mentioned crusher consists of two
primary gyratory crushing stations, so for every time point there are two energy consumption
values available. The energy consumption of the crushers are logged every 12 seconds.
The energy consumption of the crusher has a large scatter, so it is of great importance
to apply suitable statistical methods for sampling and analyzing the data. Figure 4.8 shows a
typical energy consumption diagram for one of the crusher lines. Each bar in figure 4.8
represents 12 seconds of crusher work and the values should be correlated to fragmentation of
the rock on trucks. Since the rock from a single truck takes approximately 5 minutes to be
crushed, a statistical analysis in necessary to come up with a median value for each truck.
Figure 4. 8: A sample of crusher energy consumption variations during 6 hours.
The mean values were calculated through VBA codes; the energy consumption of two
crushing lines were compared to pre-defined values to check if the crushers were in process or
idle mode; if either of them are in idle mode the other line’s energy consumption is assumed
real for that specific time point. For the time points that both crusher lines were in process, an
average of the two values are considered as true energy consumption. Over the time axis box
median values are calculated for periods of 5 minutes (similar to box-and-whisker diagrams).
It should be mentioned that the 5-minute period used in the calculations is an approximation
of the time between dumping and end of crushing, this value is obtained from observations at
the crusher and averaging the timespan.
4.3 Analysis strategy
With regard to data limitations mentioned in 4.2, a coherent approach to available data is
necessary for a consistent interrelation. Figure 4.9 shows the flow of the available sets of data
and the logical path to correlate them and bypass the data loss.
Figure 4. 9: Data-flow diagram used to correlate sets of analyzed data.
The main purpose of the diagram in figure 4.9 is to integrate all data sets into one
database of synchronized parameters. As seen, three sets of data come from the shovel; one is
the time interval of loading each bucket of the shovel, which provides a measure of digability;
second is the date and end time of loading a specific truck, which is used to designate the
affiliated photo for fragmentation analysis; and the third is the Cycle ID of the truck, this ID is
a substitute for the Truck ID, which is missing in Minestar. Truck cycle ID is then used to
extract the coordinates of the digged rock and end time of the truck cycle, which is equal to
dump time and is used to obtain the associated crusher efficiency from the crusher log.
MWD data is independent of the mentioned data sets; all available data is analyzed to
extract penetration rate and specific energy over the entire bench.
Dig coordinates for available truck cycles in Minestar, end time of loading from
shovel, and rock properties from MWD are then filtered using a VBA loop in a way that all
missing data are eliminated except the points for which all three sources have valid data.
The mentioned procedure provided a parametrically comparable set of data. Based on
this data, photos are selected and analyzed with Split-Desktop and Fragmetrics softwares. As
mentioned, Split-Desktop is used for a detailed fragmentation analysis and Fragmetrics is
only used for oversize material (boulder count).
To compare three sections of the bench, three representative areas are assumed. The
areas are selected based on the criteria of availability and validity of data, as well as
similarities in rock mass properties, which is described in 4.1.
5 TEST RESULTS
5.1 Hardness of the rock (MWD)
Since the objectives of this study do not include detailed MWD analysis, only the most
important parameters are presented and discussed. As mentioned in 4.2.1, penetration rate and
specific energy are the governing parameters in MWD analysis (Mozaffari 2007). Figure 5.1
shows the penetration rate over the test bench and specific energy is plotted in figure 5.2; the
gray area in both plots represents the region with no available data. Although variation in
penetration rate is the simplest indicator of rock mass strength, it is influenced by several
factors. Since specific energy merges all factors into one single parameter, it can be used as a
substitute for penetration rate.
Figure 5.1 shows two main zones in the bench and few hotspots indicating harder
rock; the mentioned zones are significantly less differentiated in matter of specific energy
(Fig 5.2). However, a comparison between figures 5.1 and 5.2 shows very few distinctions
regarding the hotspots. These figures can be paired to the geology map of the bench (Fig 4.2).
Figure 4.2 shows dykes of Biotite and Amphibolite Gneiss in the bench, these dykes are
approximately located at the hard rock spots in figures 5.1 and 5.2; the higher strength of
Biotite and Gneiss also confirms this supposition.
Figure 5. 1: Penetration rate of drill bit.
Figure 5. 2: Specific energy consumed for drilling over the test bench.
These figures are mainly used as the basis for the selection of representative areas for
three sections of the bench. The areas are selected in parts of the bench with acceptably
similar rock mass hardness in a way that the hard rock spots are avoided as well as the areas
close to the border of the sections, which cannot be a fair representative of the whole section.
It should be mentioned that the missing data in MWD database along with the constraints
regarding Minestar data availability did not permit any better choice of areas (Fig 5.3).
Figure 5. 3: Representative areas for three sections with respect to limited MWD and Minestar data.
The surface of the bench was laser-scanned right after the blast, the raw surface is shown in
Figure 5.4. For a better comparison the 2D-projection of this surface is shown in Figure 5.5,
which indicates the variation of swelling over the sections.
Figure 5. 4: 3D view of test bench's surface after the blast.
Figure 5. 5: Contour map of bench surface level after the blast.
q=1.02 kg/m3 q=1.17 kg/m3 q=0.91kg/m3
The significantly larger swelling of section A in figure 5.5 is compatible with
considerably higher specific charge of this section (1.17 kg/m3) compared to sections B (1.02
kg/m3) and C (0.91 kg/m3).
Although the specific charge in section B is larger than section C, no significant
alteration is observed in the amount of swelling for these two sections.
Timespan for filling each bucket by the shovel is plotted in figure 5.6 as a measure of
digability of the blasted rock. The gray area indicates the missing data.
Figure 5. 6: Digability of the blasted rock over the bench.
The highly scattered digging times in figure 5.6 do not lead to any meaningful
conclusion; no significant difference can be distinguished between the three sections of the
The high scatter can be due to variations in machinery and operators. As mentioned
before, two shovels have loaded the test bench, one of type P&H 4100C and the other of type
Bucyrus 495BII. In addition to that each operator works with a unique pace; the digging time
highly depends on the skills of the operator and the operation shift (day/night).
Since boulders are the main problem in comminution process in Aitik, extra attention has
been paid to the oversized material. Oversized material is defined as particles larger than 100
cm, which is decided according to crushers’ opening size.
A total number of 78 photos were processed to evaluate the fragmentation of the test bench.
Respectively 23, 30 and 25 photos were analyzed for representative areas of sections A, B and
Complete size distribution diagram of three sections are presented in figure 5.7; the
curves are totally based on existing data and no curve fitting method is included. As seen in
the diagram, section B is of lowest uniformity and the most uniform curve belongs to section
A. However, sections B and C are approximately identical between X80 and X100.
Figure 5. 7: Particle size distribution of the fragmented rock.
For a better comparison the values of X50, X80 and percentage of oversize material are
presented in figures 5.8 and 5.9. Although section C has the largest X50 and X80 values, the
variations are minor between three sections.
Figure 5. 8: Size of the material at 50% and 80% passing.
Section A Section B Section C
q=1.17 kg/m3 q=1.02 kg/m3 q=0.91 kg/m3
Section A has the lowest X80; the deviation of X80 from reference bench is
approximately equal for sections A and C. Nevertheless, the percentages of oversize material
(Fig 5.9) show a notably higher deviation for section A. Section C includes the largest
percentage of oversize material, but the difference from the reference (section B) is less than
3%. However, section A includes 3.92% of oversize material, which is more than 7% lower
than reference section.
Figure 5. 9: Percentage of oversize material from Split-Desktop analysis.
Although this study mostly focuses on the coarse portion of particle size distribution
(X80 and oversize material), value of X50 is the most common measure of fragmentation;
therefore the values of X50 for three sections are plotted against their corresponding specific
charges on a log-log diagram in Figure 5.10. The line fitted to three points on the diagram
leads to the conclusion that section B does not follow the common trend of specific charge
and X50 correlation. The smaller X50 in section B, which shows finer material, could be due to
geological variations in three sections. As shown in the rock hardness diagram from MWD
analysis (Figures 5.1 and 5.2), section B consisted of slightly weaker rock compared to
sections A and C, such variation may have led to finer fragmentation and smaller X50. In
Section A Section B Section C
Oversize material (Split-Desktop)
q=1.17 kg/m3 q=1.02 kg/m3 q=0.91 kg/m3
addition to that, the errors involved in drilling, charging and measurements of fragmentation
could have led to such results.
Figure 5. 10: Logarithmic diagram of X50 versus specific charge.
The graphs in figures 5.8, 5.9 and 5.10 compare the mean values of X50 and X80 for
the processed images; yet these mean values are extracted from highly scattered data sets. A
statistical analysis of the data resulted in the diagram presented in Figure 5.11, which
demonstrates the box and whisker diagram of X50 and X80 values for each section. The graph
presents the probability density of the data. Each box, marked with the first and third quartiles
and the median value in-between, shows the range that 50% of the points are set. As seen, the
boxes include a skewness factor and show the statistical dispersion of the data rather than the
normal distribution. Overlaps of the boxes, large interquartile range (IQR) of the mean values
and wide range of minimum and maximum values can be explained by the fact that X50 and
X80 values extracted from an image from a single truck cannot be a realistic representative of
the fragmentation. Since segregation of the broken rock is an inevitable phenomenon during
loading, each truck may carry more or less homogeneously distributed materials. In other
log q (kg/m3)
words, a single truck may include mostly fine materials while the next truck carries very large
Figure 5. 11: Statistical dispersion of X50 (left) and X80 (right) values.
In Figure 5.11, X50 values as low as 5 cm or as large as 105 cm exist for particle size
distribution of image analysis; similarly, X80 values vary in range of 25 cm to 145 cm. The
only way to draw a conclusion from such wide range, which is caused by segregation of the
materials, is to sample the images in a way that includes various ranges of material sizes on
trucks so the combination of several images leads to a more representative result. Therefore
one can deduce that in order to achieve realistic results, sufficient number of images should be
analyzed so the effect of segregation of the materials is eliminated.
The size distribution curves produced by Fragmetrics software will not be taken into account
in discussions due to its nonsensical results. Four curves produced by Fragmetrics are
presented in Figure 5.12 and a sample report produced by the software can be found in
Appendix I. As seen in Figure 5.12, although the delineations were manually corrected, the
curves provide very little information regarding the uniformity and size distribution of the
Figure 5. 12: Four sample size distribution curves produced by Fragmetrics software.
Photos taken by Fragmetrics camera (on P&H 4100C shovel) are only analyzed as a
means to determine the percentage of boulders larger than 100cm. A total number of 195
images (65 images per representative area) were analyzed using Fragmetrics software.
Results of Fragmetrics analysis are presented in figure 5.13. According to the diagram,
section A includes the least percentage of oversized boulders followed by section C and B
respectively. Section A’s deviation from the reference is noticeably larger than section C’s.
Figure 5. 13: Percentage of oversize material from Fragmetrics Results.
A comparison of figures 5.13 and 5.9 reveals a large difference in percentage of
oversize material between Fragmetrics and Split-Desktop Analyses. The reason might be the
low quality of Fragmetrics images, leading to a faulty estimation of fine material. The
software cannot differentiate shadows from fine material, so all the shadows and voids in
between the particles are counted as fine material. Such error led to oversize percentages
lower than actual values.
5.5 Crusher efficiency
The correlation between coordinates of the rock and energy consumption of the crusher is
shown in figure 5.14. The energy value shown in the plot is a statistical median of energy
consumption of both lines of crusher 165; the parts for which the data was not available are
marked by gray color.
Section A Section B Section C
Oversize Material (FragMetrics)
Figure 5. 14: Energy consumed to crush the rock (KWh).
According to the plot, section A had consumed the least amount of crusher energy.
The highest energy consumptions belong to an area at the border of sections B and C. A
meaningful correlation can be observed between figure 5.14 and MWD results (figures 5.1
and 5.2); the area close to MWD hotspots, which indicate harder rock, had consumed higher
6 DISCUSSION AND CONCLUSIONS
In order to deduce practical conclusions, an overall comparison of the results is required as
well as an estimation of economic efficiencies of alternatives. Table 5.1 summarizes the most
important results for each section of the bench. The drilling depth in calculation of specific
drilling is assumed equal to planned depth of 17.5 meters for all holes. Mean values are
calculated at 95% confidence of the data sets. The standard deviations (STDV) are also
mentioned to provide a measure of the data scatter.
Table 6.1: A summary of the test results.
Section A B C
Burden (m) 6 7 7
Spacing (m) 9 9 10
Height 15 15 15
Specific Drilling (m/m3) 0.0216 0.0185 0.0167
Specific Charge (kg/m3) 1.17 1.02 0.91
Mean value 7.3 4.8 4.4
STDV 1.25 1.33 0.71
Mean value 33 28 41
STDV 12.2 15.6 14.6
Mean value 63 70 80
STDV 10.2 11.0 15.3
Percent Oversize (%)
Mean value 4 12 14
STDV 0.86 1.92 1.81
Crusher Energy Consumption (KWh)
Mean value 280 380 390
STDV 62.7 70.3 66.5
As section B is the currently used design in Aitik mine, it is assumed as the reference
to compare two other sections; in order to make an unbiased comparison between the factors
mentioned in Table 6.1, percentages of deviation from reference is plotted in Figure 6.1.
Figure 6. 15: Percent deviation from reference section.
Specific Drilling and specific charge are the main measures of cost estimation. As
seen in Figure 6.1, section A had about 15% higher cost of drilling and charging compared to
section B. Same costs are -10% for section C, indicating a large thrift in total cost.
High specific charge of section A resulted in 50% more swelling compared to
reference section. However, section C had only 8% less swelling than section B.
The deviation of X80 in sections A and C are +14 and -10% respectively, which is
fairly uniform. But the percentage of oversize material shows very different deviations;
section A included 60% less boulders compared to section B; but section C included only
16% more oversize material.
Crusher energy consumption is also differently deviated for sections A and C; the
crushing of the rock from section A consumed 25% less energy than the rock from section B.
Percent deviation from reference section
In contrary, energy consumption of the crusher was only 2% higher than the reference for
section C. In other words, sections B and C consumed fairly equal amounts of energy in the
crusher. Comparing MWD analysis results to crusher energy consumption plot (Figure 5.14)
reveals a meaningful correlation between crusher efficiency and hardness of the rock; the
hotspots of MWD plot, indicating harder rock, consumed significantly larger amounts of
crushing energy. To summarize, the following conclusions can be drawn from the results:
- Section A, with a 6x9 m drilling pattern and 1.17 kg/m3 of specific charge, provides a
smaller X80 and lower percentage of oversize material; it also reduces the energy
consumption of the crusher by around 25%. However, X50 shows a 13% increase
compared to reference section. The costs of drilling and charging also increase by 15%.
- Section C, with 7x10 m drilling pattern and 0.91 kg/m3 specific charge, produced
coarser fragmentation and more boulders; in addition to that, visual observations
confirmed that the size of the boulders were significantly larger than the reference.X50
and X80 show 14% and 41% increases respectively. The percentage of oversize
materials also increased by 17%. The fragmented rock consumed only 3% more
crushing energy compared to the reference section, which is almost equal to reference
section energy consumption. Furthermore, the drilling and charging costs of section C
were 10% lower than the reference costs.
7 RECOMMENDATIONS AND FURTHER STUDIES
Two drilling patterns have been tested, evaluated and compared to the currently used pattern;
the test results are in accord with the theoretical correlations of specific charge and
In order to put the results into practice, comprehensive economic analyses as well as
another comparative test seem necessary. If the post-blast benefits of 1.17 kg/m3 of specific
charge in 6x9 m drill pattern overcome the additional costs of extra drilling and charging, a
confirmative test should approve the results prior to application of the new drilling pattern.
Since the deviations from reference are relatively low for section C (Fig. 6.1), the
7x10 m pattern and 0.91 kg/m3 of specific charge can still be considered as an alternative
option. The increase in X80 and oversize materials are about 15%, but the energy consumption
of the crusher does not increase significantly and the pattern saves 10% of drilling and
The study suggests advantageous practices of MWD database to predict spots of hard
rock and low crusher efficiency. This data can be used to modify the drill and blast design of
benches to reach uniform fragmentation throughout the whole area. The meaningful
correlation between hard rock indicators on MWD maps and significantly larger crushing
energies can also be used to eliminate the effect of geological uncertainties in future tests
Minestar system is another potential for improvements; a well-functioning logging
system opens the way for continuous surveys of mining process. Minestar logs act as a link
between different sections of the mine, so in order to improve the overall efficiency of the
mine it is critical for Minestar system to fully function.
Fragmetrics system is also an advantageous means for continuous monitoring of
fragmentation. However, low image quality and faulty analysis algorithms do not let this
system to be used in full capacity. A high resolution camera equipped with an anti-vibration
system, together with more advanced image processing software, can provide more realistic
results in regard of continuous fragmentation measurements. Finally, following operations are
suggested for future:
- A financial analysis of tested patterns; a confirmation test should be defined if either of
them leads to higher benefits.
- Efforts to improve the function of Minestar and mine network in order to collect as
much data as possible.
- Efforts to improve the function of Fragmetrics system as a key in fragmentation
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APPENDIX I: Fragmentation analysis of FragMetrics software.
Region: S1_210_13 Shovel: 1152 Number of bucket images processed: 462 Date range: May 9, 2012 12:22:16 AM GMT+2:00 ~ May 20, 2012 5:31:56 AM GMT+2:00 Duration: 11 days 5 hours 9 minutes Calibration bucket width: 460 cm Report date: May 20, 2012 1:31:26 PM GMT+2:00 History view of P10 to P100
Schumann distribution modulus = 1.119 Schumann size modulus = 2.8 cm Table 2 Fragmentation Target Parameters
Oversize 100 cm Undersize 5 cm
Table 3 Fragmentation Results
Percent Oversize 3.1 % Percent in Range 27.1% Percent Undersize 69.8%
Generated by: Fragmetrics™ - Tablet
iii APPENDIX II: An introduction to Air-decking in Aitik.
The utilization of air-decks in production blasting is a fairly new method in mining
industry, although the concept has been studied and practiced since 1986. Melnikov was the
first to introduce air gaps inside the blast hole and most of the early research on this topic has
been conducted in the former Soviet Union (Lu and Hustrulid 2010).
The main concept of air-decking consists of decreasing the amount of explosive and
improving the blast-induced fragmentation by means of introducing air gaps in the explosive
column. Theoretical studies and field experiments by Melnikov and Marchenko(1971),
Chiappetta and Mammele(1987), Bussey and Borg(1995) and Jhanwar et al. (1999) show a
decrease in mean fragment size, an increase in the uniformity of fragmentation, and also a
decrease in explosive consumption by 10-30%. Laboratory experiments carried out by
Fourney et al. (1981) regarding air decking in thick plexiglass blocks revealed that air
decking increases the effect of shockwave on the material by a factor of 2 to 5. Lu and
Hustrulid (2003) reviewed the theory of airdecking and provided guidelines regarding the
application of this method.
In spite of these studies, the mechanism of airdecking has still not been fully
understood and utilization of this method does not always improve blasting results.
Air-decking technique comprises the use of one or several air gaps in the explosive
column in order to optimize the fragmentation and reduce the explosive consumption. The
theory proposed by Melnikov and Marchenko (1971) hypothesizes that the air gap is a means
of shockwave reflection within the borehole. The air-deck acts similar to a cushion and
produces a series of aftershocks that extend the network of microfractures in the rock. The
aftershocks are produced by three main pressure fronts: shock front, pressure front due to
iv formation of explosion products behind the detonation front and reflected waves from the end
of explosive column. Although the air-deck causes a reduction in borehole pressure, the
repeated loading of the rock by a series of aftershocks prolongs the action time of the
shockwave and results in improved breakage (Jhanwar et. al 1999).
A series of tests by Fourney et al. (1981) in Plexiglass models supported Melnikov’s
theory. It was observed that the shockwave reflects back from the base of the stemming
column and reinforces the stress field. This process is repeated several times; therefor the
duration of the shockwave action is increased by a factor of 2-5. This mechanism leads to a
larger volume of radially fractured material rather than heavy breakage in the area adjacent to
the charge column, see Figure I.1.
Figure I. 1: Development of crack network in Plexiglass under the influence of an air-decked explosive column (after Fourney et al. 1981).
Despite the confirmative results from field and lab tests, some important technical
problems are still unsolved. The location of the air deck in the blast hole and the length of the
air column are two of the main questions to which there are different answers proposed.
Moxon et al. (1993) showed that as the length of the air-deck is increased the fragmentation
v becomes finer relative to that of a full-column charge. The reduction however is relatively
small until a critical length is exceeded. The critical length depends on the strength and
structure of the rock mass. From the model tests, a critical air-deck length of 30–35% of the
original explosive column was determined. They concluded that a mid-column air-deck has a
larger effect on fragmentation than that of the top or bottom air-deck. Liu and Katsabanis
(1996) and Katsabanis (2001) found that there exists a minimum air deck length for the
technique to be beneficial; they also found that variations to the top air deck, such as bottom
and mid-column air decks do not make significant improvements in production blasting.
Recent investigations by Hustrulid et al. (2003) also show that in case of top air decks, there
is a minimum limit for the air-deck length to be effective on the fragmentation. The length of
the air-deck is the most important parameter, so this limit has been determined empirically by
several tests and it is presented as the “Air-decking ratio”, which is the ratio between the
length of the air deck and the total length of the explosive column:
where Ra is the air-decking ratio, La is the length of the air-deck and Le is the length of the
explosive in the column. The corresponding value for R, according to Hustrulid et. al (2003),
374.0164.0 ≤≤ R (I.2)
For the current case in Aitik, the practical constraints do not allow a ratio as high as
0.3, so the following designs are suggested based on practical applicability of them in Aitik.
As seen in Figure I.2, a total length of 18m has been assumed for the boreholes. The
length of the air-deck is suggested in accordance to rules of thumb and applicability in the
site. The length of the deck is at minimum 2.5 meters that includes a safety margin for some
immerging of the barrier into the emulsion.
Air-decking ratio and amounts of reduction in explosive are mentioned in table 1 to
provide a comparison between the options. The price of emulsion explosive and the diameter
of the borehole are assumed 5 SEK/litre and 318mm respectively4.
Figure I. 2: Suggestions for air-decking in Aitik; a)original design, b)2.5m air-deck, no charge reduction, c) 3m air-deck, reduced charge, d) 2.5m air-deck, reduced charge.
Tabe I. 1: A comparison of the designs and cost reductions per hole.
Type Charge column (m)
Air column (m)
(SEK) a (original) 12.5 0 5.5 0.00 0 0.0
b 12.5 2.5 3 0.17 0 0.0 c 11 3 4 0.21 120.6 603.2 d 11.5 2.5 4 0.18 80.4 402.1
Since an extra cost will be added for the air-decking instrumentation, the c and d
designs are favorable, because of their lower explosive cost. Between c and d, the latter has
also the advantage of smaller reduction in amount of charge and an acceptable stemming
4 The price is a rough fictitious assumption.
vii iv. Implementation of air-deck
The air gap in the blast hole is most commonly implemented through usage of
balloons or gasbags. The gasbag or balloon is lowered to the desired depth in the blast hole
and inflated; upon inflation, the air pressure inside the balloon fixes it inside the hole and
depending on the location of the air-deck, stemming or explosive is filled on top of the
Different types of gasbags and balloons are available on the market, which can be
categorized into two main groups: Chemically inflated and mechanically inflated. Chemical
gasbags are inflated by means of chemical reactions of the material inside the bag; the
reaction is either between two substances that are mixed when a button is pressed, or by a
pressurized gas capsule. These gasbags are easy to implement and fairly cheap; they also take
short time to install in production holes. Mechanically inflated gasbags are more like regular
balloons modified to withstand high pressure of blast hole. These balloons are inflated
through an air compressor, which can be easily installed on any truck. These balloons take
comparatively shorter time for installation (10-20 seconds) and usually have a smaller chance
of failure. The installation method for each group varies depending on the gasbags’ type and
manufacture, see Figure I.3.
Figure I. 3: Left: Samples of commercially available balloons (Left) and gasbags (Right).
v. Air-decking and Aitik conditions
Aitik mine is located in northern Sweden with long and harsh winters; the temperature
in the winter can reach -40°C. The extremely low temperatures can cause malfunctions in
chemical gasbags or pressurized balloons. In addition to that, high level of groundwater
causes the blast holes become filled with water almost instantaneously after drilling; lowering
an inflating gasbag 3-4 meters deep into a water-filled hole is not a fast and easy task for
production blasting. Previous tests of capsule-inflated gasbags were not successful due to low
air pressure in cold weather; the technicians also faced difficulties in implementation process
of the gasbags in the water-filled blast holes.
Gasbags and balloons are usually lowered into the desired depth by a rope. In case of
Aitik mine, a long piece of wood is used to force the bag into the water-filled holes. In a
future tryout a modified type of capsule-inflated gasbags are going to be tested; the capsules
are expected to function normally in cold weather, but the mechanism of inflation will still be
problematic in water-filled areas of the mine. These gasbags have a diameter less than the
ix desired one; the capsule inside the gasbag should be triggered on surface and only 20-30
seconds after triggering the gasbag will pop to the desired diameter. For the case of water-
filled holes, it might be challenging to lower a half-inflated gasbag into the blast hole during
such short interval.
Mechanically inflated balloons appear to be a better choice than gasbags. They are
usually lowered into the hole by a hose, which is designed to come off when the balloon is
inflated. The balloons are lowered in deflated estate, so the technician will face less difficulty
in lowering the balloon into the water-filled holes. The balloon is then inflated by means of
an air compressor on the truck, so the inflation process takes less time than capsule-inflated
gasbags and the technician should only hold the balloon in place for a few seconds until it is
fixed in the hole.
It should also be mentioned that the material used in manufacture of balloons are
supposed to be more suitable for high air-pressure and heavy load of stemming on top of the
balloon. Some stereotypes of capsule-inflated gasbags were tested on the surface and there
were cases of rupture by air pressure.
vi. Conclusion and recommendations
Although the mechanism has not been fully understood, air-decking is a useful
technique to improve the upper portion of fragmentation in production blasting. The field
tests and successful cases of application in other mines support the theory behind it. But air-
decking is not beneficial for all mines. In case of Aitik mine, the first question is whether this
technique can make an improvement in fragmentation or not. Secondly, the technique should
be adapted to the harsh conditions in Aitik, which had led to unsuccessful tryouts earlier.
Between two main types of commercially available air-decking gears, the balloons
appear to fulfill more requirements than capsule-inflated gasbags, but it is recommended to
test both types and observe the results. The tests should include both cold weather and water-
x filled holes conditions. It is also recommended to test them as a simulated production blast i.e.
long time stay in the hole before the blast, etc.
Depending on the results of the tests, a full-scale blast can be air-decked and blasted.
Careful monitoring of the fragmentation and comparing it to normally blasted benches may
reveal whether the technique will be beneficial in Aitik or not.
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