Drilling and Blasting-MM 321

Ms Dhelda Mfanga
Room Q 204
 Part One- Blasting

Topic 1- EXPLOSIVES
INTRODUCTION
 The use of explosives in mining and construction applications dates back to
1627
 From 1627 through 1865, the explosive used was black powder.
 In 1865, Nobel invented nitroglycerin dynamite in Sweden. He invented also
gelatin dynamites in 1866.These new products were more energetic than
black powder.
 In the mid-1950's' a new product appeared which was called ANFO,
ammonium nitrate and fuel oil. This explosive was more economical to use
than dynamite.
 Since then, ANFO has become the workhorse of the industry and
approximately 80% of all explosives used in the world.
 Other new explosive products appeared on the scene in the 1960's and
1970's. such as slurries or water gels.
 In the late 1970's' a modification of the water gels called emulsions appeared
on the scene.
 Therefore, commercial explosives fall into three major generic categories,
dynamites, dry blasting agents and wet blasting agents (slurries or water gels
or emulsions).
What is Explosive?

 It is unstable chemical compound or a mechanical mixture when properly

INITIATED by heat, impact, friction or their combination it undergoes rapid
disintegration called EXPLOSION.
 Upon ignition there is a rapid release of heat (up to 45000 C) and large
quantities of high-pressure gases (up to 250,000 bar) that expand rapidly
create a shock wave called DETONATION.
 The resulting detonation travels through the un-reacted explosive and the
surrounding rock at a sonic speed ranging from 1500 to 9000 m/s: from 1500
to 2500m/s for low explosives and greater than 2500 m/s for high explosives;
 The energy released by the detonation of explosives produces four basic
effects; (a) rock fragmentation; (b) rock displacement; (c) ground vibration;
and (d) air blast
 The main elements of explosives are Carbon (C), Hydrogen (H2), Nitrogen
(N2), Oxygen (O2) and Aluminum (Al).
THE PROCESS OF EXPLOSIVE DETONATION

 The self-sustained shock-wave produced by a chemical reaction in a

gaseous medium was described by D.L. Chapman and E. Jouquet as a
space.
 This space of negligible thickness is bounded by two infinite planes - on one
side of the wave is the un-reacted explosive and on the other, the exploded
gases, as shown below.

 There are three distinct zones: (a) the undisturbed medium ahead of the
shock wave; (b) reaction zone and (c) zone of expanding gases. This
condition for stability exists at hypothetical X, which is commonly referred to
as the Chapman-Jouquet (C-J) plane
C-J Plane
 The process of explosive detonation is illustrated on figure below as idealized
shock wave propagation through a cylindrical explosive
 The steady-state chemical reaction takes place behind the shock front within
the reaction zone.
 At the end of this zone, a nonsteady-state region exists.
 The C-J plane is seen as the boundary between the steady and non-steady
state, where the reactions are considered complete.
 This is also the plane where all the thermodynamic properties are calculated.
These are: pressure, velocity, temperature, internal energy or heat of
formation and density
THEORY OF EXPLOSIVES IN BREAKING THE ROCK

A general theory of explosives in rock breakage is that:

 The detonation of the explosives charge causes a high-
velocity shock wave and a tremendous release of gas.
 The shock wave cracks and crushes the rock near the
explosives and creates thousands of cracks in the rock.
 These cracks are then filled with the expanding gases.
 The gases continue to fill and expand the cracks until the gas
pressure is too weak to expand the cracks any further, or are
vented from the rock.
CLASSIFICATION OF EXPLOSIVES

Low energy Explosives

 Low explosives deflagrate rather than detonate. Their reaction
velocities are from 1500 to 2500m/s.
 They are characterized by a pressure pulse of lower amplitude and
longer duration
 They don’t produce shock energy. They produce work only from gas
expansion.
 A very typical example of a low explosive would be black powder
High energy Explosives
 High energy explosives are characterized by the intense shattering
effect, known as brisance upon detonation
 They detonates with a reaction velocity over 2500 m/s.
 The distance between the shock wave front and the C-J plane is
very short and results in a pressure pulse of high amplitude and
short duration.
 In a high explosives, there are two distinct and separate pressures;
shock pressure and gas pressure.
 Shock pressure is estimated to account for only 10% to 15% of the total
available useful work energy in the explosion.
 The gas pressure accounts for 85% to 90% of the useful work energy.
 During a detonation in high explosives, the shock energy at the reaction front
travels through the explosive before the gas energy is released.
 This shock energy, normally is of higher pressure than the gas pressure
 The gas energy in high explosives is much greater than the gas energy
released in low explosive

Pressure Profiles for Low and High Explosives

MECHANISMS OF ROCK BREAKAGE

PHASE 1: Explosive detonates

• During the detonation process, the explosives
are quickly converted into a high temperature
and high pressure gas

PHASE 2:1: Shock wave damage

• Immediately around the blasthole, the
high detonation pressures propagate a
shock wave into the rockmass as a
result a zone of between 2 to 3 charge
diameters is crushed.
• Shock wave is then reduced to strain
pulse.
PHASE 2:2: Tensile stress damage
As the strain pulse propagates through the
rockmass at a rate equal to the P-wave
(seismic velocity), it compresses the rock
radially which results in tangential stress at
right angles to the compression wave front.

• If the tangential stress is greater than

the tensile strength of the rock
fractures are created (from typically 20
to 30 charge diameters) that extend in
all directions( radial cracks)

• When the strain pulse reaches a

rock/air interface (such as a open joint,
free face) the pulse is reflected back
tend to open those radial cracks.
PHASE 3:Crack extension
• This phase starts at the same time as P1
& P2 but takes place at a much slower
rate.
• The high pressure gases wedge into
existing cracks and cause them to
expand.
• The gases continue to fill and expand the
cracks until the gas pressure is too weak
to expand the cracks any further

PHASE 4:Pressure Heaving

• The last part of the fragmentation process
occurs as gas pressure bends and breaks
the rockmass (flexural rupture) toward the
path of least resistance
THE CHARACTERISTICS AND PROPERTIES OF EXPLOSIVES
1) Density
 It is the weight per unit volume of the explosive – specific gravity, kg/m3. The
specific gravity of commercial explosives ranges from 0.6 to 1.7 g/cc.
 For difficult blasting conditions or where fine fragmentation is required, a
dense explosive is usually necessary.
 In easily fragmented rock or where fine fragmentation is not needed, a low-
density explosive will often suffice.
 For water filled borehole, explosives whose density is greater than 1 kg/ m3
should be used.
Borehole Loading Density
- It is the weight of explosive which could be loaded in a unit length of
borehole, kg/m. It is determined by the explosive density (ρ) and borehole
diameter (D):

LD = π [D/2]2 ρ, kg/m
2) Water Resistance
 An explosive's water resistance is a measure of its ability to withstand
exposure to water without deteriorating or losing sensitivity.
 Explosives vary widely in water resistance – ANFO - no water resistance,
emulsions, water gels - good water resistance.
 Higher-density explosives have fair to excellent water resistance, whereas
low-density explosives and blasting agents have little or none.
 Brown-orange nitrogen oxide fumes from blast indicate inefficient detonation
which might have been caused by wet explosives.
3) Shelf life
 Explosives deteriorate and shelf life is particularly affected by both
climate and magazine conditions.
 For most explosives products, a shelf life of one year is
recommended, although satisfactory performance can be expected
from most products two, three, and even four years later.
 Explosive manufacturers specify the storage properties or shelf life
of their products, based on normal magazine conditions (ICI 1997).
4) Fume class
 Characterizes the noxious gases (NO, CO, H2S, SO2), etc,
generated from the disintegration of explosive. Consideration of
fume class is important in the selection of the type of explosive.
 For open work, fumes are not usually an important factor, In
confined spaces(underground mine), however, the fume rating of an
explosive is important. In any case, the blaster should ensure that
everyone stays away from fumes generated in a shot.
 Factors increasing toxic fume generation: improper priming, lack of
confinement, water, improper explosive composition, improper
timing, improper loading techniques and adverse reaction with rock
 Figure 2. Excessive fume production
5) Sensitivity
 Measure of the ease of initiation of explosive. Also it is the
resistance of explosive to accidental detonation.
 It is established through the “Cap sensitivity test” involving measure
of the minimum energy required to detonate the explosive.
 Explosives are classified as cap sensitive or non sensitive based on
the No. 8 standard cap. Example: high explosive (1.1D) - sensitive to
a No. 8 strength blasting cap and blasting agent (1.5D) - not
sensitive to a No.8 strength blasting cap - requires booster
6) Strength
 It is a measure of the ability of explosive to break rock /performance
potential.
 It is determined by the detonation velocity and density of explosive,
the volume of the gaseous products of explosion and the heat of
explosive reaction
 Explosive strength is measured as the absolute bulk strength (ABS)
in Joules/m3 or the absolute weight strength (AWS) in Joules/kg or
Relative bulk strength (RBS)
7) Critical Diameter
 It is the minimum borehole diameter (Dmin) in which a particular
explosive compound will detonate reliably.
 At diameters below the critical diameter, explosive reactions are
incomplete and energies lower than the available capacities are
liberated.
 Above the critical diameter, the detonation of explosive travels at the
maximum- steady state velocity and all thermodynamic parameters
are at their maximum values.

Figure 3:Critical borehole diameter of explosives

8) Temperature Effect
 Low temperature affects the stability and performance of explosive.
 The sensitivity and detonation velocity of certain water based
explosives is hampered.
 Dynamites could become dangerously unstable in below freezing
temperatures.

9) Detonation and borehole Pressure

 Pressure produced in primary reaction zone of explosive
Pd(mPa)= 250 x explosive density x VOD ^2
 A high detonation pressure is necessary when blasting hard, dense
rock. In softer rock, a lower pressure is sufficient.
 Borehole pressure is almost 30 % less than the theoretical
detonation pressure
10) Detonation velocity (VOD)
 It is the speed at which detonation shock wave front travels through
a column of explosive charge, m/s.
 For high explosives (HE), the strength of explosive is directly
proportional to its detonation velocity.
 The factors influencing on the detonation velocity of explosive
charge are the field borehole loading conditions such as borehole
diameter, loading density, confinement, the presence of water, etc.

EXPLOSIVE DETONATION
11) Oxygen Balance (OB)
 OB is the explosive content of excess or deficit oxygen compared to the
content just sufficient for the complete oxidation of fuel elements within the
explosive.
 The basic elements or ingredients which directly produce work in blasting are
those elements which form gases when they react, such as carbon,
hydrogen, oxygen, and nitrogen.
 If only the ideal reactions occur from the carbon, hydrogen, oxygen, and
nitrogen, there is no oxygen left over or any additional oxygen needed. The
explosive is oxygen balanced and produces the maximum amount of energy.
 If two ingredients are mixed together, such as ammonium nitrate and fuel oil,
and an excess amount of fuel oil is put into the mixture, the explosive
reaction is said to be oxygen negative.
 If too little fuel is added to a mixture of ammonium nitrate and fuel oil, then
the mixture has excess oxygen which cannot react with carbon or hydrogen.
This is called an oxygen positive reaction.
 The maximum detonation energy is obtained from explosives formulated with
zero oxygen balance, whose explosion products are vaporous H2O, CO2
and N2.
 The energy is reduced because other ideal gases liberate heat when they
form, nitrogen oxides absorb heat in order for them to form.
Carbon-Oxygen Ideal Reaction Nitrogen-Nitrogen Ideal Reaction

Non-Ideal Carbon-Oxygen Reaction Non-Ideal Nitrogen-Oxygen Reaction

Identification of Problem Mixtures
THE COMPONENTS OF COMMERCIAL EXPLOSIVES
I. Explosive bases
 An explosive base is a solid or a liquid which, upon application or heat or
shock, breaks down very rapidly into gaseous products, with an
accompanying release of heat energy. Nitroglycerine an example
II. Combustibles
 A combustible combines with excess oxygen in an explosive to achieve
oxygen balance, to prevent the formation of nitrous oxides (toxic fumes),
and to lower the heat of the explosion
III. Oxygen carriers (oxidizers)
 Oxygen carriers assure complete oxidation of the carbon in the explosive
mixture, which inhibits the formation of carbon monoxide. The oxygen
carriers assist in preventing a lowering of the exploding temperature. A
lower heat of explosion means a lower energy output and thereby less
efficient blasting.
IV. Antacids
 Antacids are added to an explosive compound to increase its long term
storage life, and to reduce the acidic value of the explosive base,
particularly nitroglycerin (NG
V. Absorbents
 Absorbents are used in dynamite to hold the explosive base from
exudation, seepage, and settle-ment to the bottom of the cartridge or
container. Sawdust, rice hulls, nut shells, and wood pulp are often used as
absorbents
THE COMPONENTS OF COMMERCIAL EXPLOSIVES

VI. Fuels
 Fuel oil, carbon, granular or powdery aluminium, TNT,
black powder or any carbonaceous material which
produces heat. Many of them are also sensitizers and
absorbents.
VII. Stabilizers
 Flame retardants, gelatins, densifiers, emulsifying agents
and thickeners;
VIII. Sensitizer
 Provides the heat source (‘hot spot’) to drive the
chemical reaction. Sensitizers are generally small air
bubbles or pockets within the explosive
VIII. Antifreeze
 Antifreeze is used to lower the freezing point of the
explosive
THE COMPONENTS OF COMMERCIAL EXPLOSIVES
 CATEGORIES OF COMMERCIAL EXPLOSIVES

 The speed of detonation is one of the parameters used to classify an

explosive
I. THE LOW ENERGY (LE)
 Black powder, blasting powder, Judson powder, etc.
 Leave almost ½ of their volume as solid residue after ignition.
 Generate large volumes of gases at slow rate, exerting a pushing
effect on the surroundings.
 Don’t exhibit shock wave. Burn with a rapid rate called deflagration
 BLACK POWDER: 72 % NaNO3 + 16 % Charcoal + 12 % Sulfur. It
is cheaper than blasting powder but good water absorbent;
 BLASTING POWDER: 72 % KNO3 + 16 % Charcoal + 12 % Sulfur.
It is more expensive but having better keeping qualities.
 JUDSON POWDER: It is black powder in which hard, porous grains
are coated with nitroglycerin of content varying from 5 to 20 %.
Stronger than the black and blasting powders.
II. HIGH ENERGY (HE)
 Those used for military such as TNT, PETN, RDX, etc
 Those used commercially such as dynamites, blasting
gelatins, ammonium nitrate explosives, etc. completely
gasified when properly initiated
 Exert strong shattering effect on the surroundings because of
their rapid gases liberation process upon detonation
 DETONATION: It is the process of shock wave propagation
through explosive, accompanied by the energy needed to
maintain propagation in a steady state.
FORMULATION OF COMMERCIAL EXPLOSIVES
 If cotton, starch or glycerin is treated with nitric acid, product is nitro-cellulose
(guncotton), nitro-starch and nitro-glycerin respectively;
 Nitro-glycerin and nitro-starch are also mixed with other substances to make
dynamites and other high explosives.
Nitroglycerin (NG)
History
 Discovered by Sobrero in 1847 in Paris but Alfred Nobel was the first to
manufacture it commercially.
 Alfred Nobel and his father built a NG small factory in 1861 in Sweden
 Loading and transporting nitroglycerin was dangerous (liquid poured in to
holes and ignited with various types of black powder igniters.
 NG proved to be very dangerous and resulted in the death of many people
including his brother Emil
 Nobel discovers dynamites by accident (Dynamite is derived from the word
Dynamis, meaning power):
 When ‘blasting oil’, NG spilled into kieselguhr (NG was packed in it), Nobels
saw that the kieselguhr abosrobed about 3 times its weight of NG.
 Nobel began to sell the 75/25 NG/kies., the first of the dynamites.
 Eventually went to wood pulps which increased the energy output of
the NG.
 This development allowed the relatively safe transport transportation
of NG.

Properties of NG
 Not soluble in water, Poisonous, Liquid and dangerous to handle.

 Highly sensitive, highly powerful.

 Due to its high sensitivity and liquid form, nitroglycerin is dangerous
for transportation and in its present form unsuitable for usage in the
Mining Industry.
 Chemical formula: C3H5 (NO3)3; density: 1.6; oxygen balance:
positive; commercial use: used as an explosive base for commercial
explosives.
C3H5O9N3 3CO2 + 2.5H2O +1.5N2 + 1503 cal/gm
Fertilizer Grade Ammonium Nitrate (AN) + Fuel Explosives
 In 1955, Ammonium nitrate/fuel oil (ANFO) was discovered by
Dupont to greatly increase the energy output of AN prills.
 Fertilizer grade AN is made up of prills from which the last 5
% H2O is evaporated after the prills are solidified producing
porous prills with more surface area per unit weight than solid
AN.
 High porosity makes AN easily detonated.
NH4NO3 + 0.5C N2 + 2H2O + 0.5 CO2 + 929,
cal./gm.
NH4NO3 + 1/3CH2 N2 + 1/3H2O + 1/3 CO2 + 927,
cal./gm.
 The AN + FUEL explosives are classified as blasting agents
rather than explosives and having the following properties:
Low density; Soluble in water; Low sensitivity and Cheaper
than nitroglycerin.
BLASTING AGENT

A blasting agent is an explosive that:

 Comprises ingredients that by themselves are non-explosive
 Can only be detonated by a high explosive charge placed within it
and not by a detonator.
 All blasting agents contain the following essential components

Oxidizer A chemical that provides oxygen for the reaction.

Typical oxidizers are ammonium nitrate and calcium
nitrate.
Fuel A chemical that reacts with oxygen to produce heat.
Common fuels include fuel oil and aluminium.
Sensitizer Provides the heat source (‘hot spot’) to drive the
chemical reaction of oxidizer and fuel. Sensitizers are
generally small air bubbles or pockets within the
explosive.
 TYPES OF COMMERCIAL EXPLOSIVES

- STRAIGHT DYNAMITE
- GRANULAR
- HIGH DENSITY EXTRA DYNAMITE
DYNAMITE
- LOW DENSITY EXTRA DYNAMITE
DYNAMITE
- GELATIN - STRAIGHT GELATIN DYNAMITE
DYNAMITE - AMMONIA GELATIN DYNAMITE
- SEMIGELATIN DYNAMITE

- WATER GELS (SLURRIES)

- BULK
WET BLASTING - EMULSIONS
- CARTRIDGED
AGENTS - HEAVY ANFO
more than 5% water by weight

DRY BLASTING - BULK ANFO

AGENTS (ANFO) - CARTRIDGED ANFO
DYNAMITES
 They are nitroglycerin - nitrostarch mix.
 Basic types are granular dynamite, gelatin dynamite, semi-gelatins dynamite.
 Gelatins and semi-gelatins: NG + nitro-cotton (Gel structure whose
consistency is controlled by the % of nitro-cellulose)
 Dynamites are the most sensitive of all the generic classes of explosives
used today
 Dynamites are packed in cylindrical cartridges of diameter 22 mm and length
varying from 203 to 610 mm.

STRAIGHT DYNAMITE
 NG, NaNO3, carbonaceous fuels, sulfur and antacids;
 Absorbents + Combustibles: wood pulp + flour;
 Strength: (40, 50, 60) % NG;
 Characteristics: High detonation velocity & brisance, low flame temperature,
good water resistance & shock sensitivity, poor fume qualities.
EXTRA DYNAMITE
 Granular mix of NG and AN +SN;
 AN – dynamite varieties: high density and detonation velocity, (20 – 60) %
NG;
 Characteristics: Poor water resistance, low explosion temperature and
density, with high or low detonation velocity;
 Strength: 65 % NG.
GELATIN DYNAMITE
 Straight gelatin or AN (extra) gelatin, basically blasting gels
 Gelatins are used as boosters and primers;
 Characteristics: High water resistance and least fumes liberation.
SEMIGELATINS
 Same as AMMONIA GELATIN except that more of the nitroglycerin,
nitrocellulose mixture and sodium nitrate is replaced by ammonium nitrate.
 Strength: 65 %NG;
 Characteristics: Higher water resistance than AN dynamites but not as strong
as gelatin dynamites;
 Used as primers and boosters.
DBA- BULK ANFO
 A dry agent is a granular, free-running mix of a solid oxidizer (usually
AN),prilled into porous pellets onto which a liquid fuel oil or propellant is
absorbed.
 Bulk ANFO (Ammonium Nitrate and Fuel Oil) is by far the most commonly
used explosive
 Improper mixes produce lower explosive energy and noxious gases
 The proper mix is oxygen balanced and occurs with a mix of 94.3% AN and
5.7% FO by weight
 The properties of dry blasting agents vary significantly with borehole
diameter, density, confinement, particle size, water conditions, and size of
primer used for initiation.
 Density: 0.75 – 0.95;
 Detonation velocity > 4500 m/s in holes of diameter > 381 mm;
 Critical diameter varies from 51 – 102 mm.
 Inexpensive, simple to manufacture
 Not waterproof
 Can include various amounts of aluminum for extra energy
VOD vs. borehole diameter for selected industrial explosives
Bulk ANFO truck
EMULSIONS
 Consists of oxidizers dissolved in water surrounded by a fuel - fine
particle size
 Relatively expensive compared to ANFO
 Very water resistant in full concentration
 Plant or truck mixed
 Typical density range of 1.15 to 1.45 g/cc
 Typically high detonation velocity and bulk strength

Emulsion pump truck

WATER GELLS OR SLURRIES
 Colloidal suspension of solid AN particles in an AN liquid solution that
is gelled
 Have coarse particle sizes
 High density and bulk strength,
 Relatively expensive compared to ANFO
 Very water resistant in full concentration
 Plant or truck mixed
 Coarse particle size reduces detonation velocity
HEAVY ANFO
 Consists of prills ANFO mixed with emulsions or watergels in
varying percentages
 Developed in an attempt to increase the bulk density of ANFO
 Full range of cost between those of ANFO and emulsion;
 Varying degrees of water resistance
 Plant or truck mixed
 Bulk trucks are designed to blend components prior to loading.
Heavy ANFO
BAGGED PRODUCTS
 Can make ANFO water resistant
 Reduced pounds of explosive per foot of borehole unless package is
split
 If the package is not split to achieve good coupling the borehole
pressure and energy output per foot are substantially reduced
 Should be lowered, not dropped, down dry boreholes

Bagged ANFO
PRIMERS AND BOOSTERS
 A primer charge is an explosive ignited by an initiator, which, in turn, initiates
a noncap-sensitive explosive or blasting agent.
 A primer contains cap-sensitive high explosive ingredients.
 Often cartridges of dynamites, highly sensitized slurries, or emulsions are
used with blasting caps or detonating cord.

Primer (cast booster and detonator)

• Boosters are highly sensitized explosives or blasting agents,
used either in packages of weights greater than those used
for primers.
• Boosters are placed within the explosive column where
additional breaking energy is required
• Boosters are often used near the bottom of the blasthole at
the toe level as an additional charge for excessive toe burden
distances.

Dynamite and cast booster

PERMISSIBLES
 Made from AN or NG base which produces short lived
detonation flames so that they don’t ignite methane or coal
dust;
 Sodium chloride or water is added in the mixture to absorb
some of the heat generated or carbon to cause the
incomplete combustion (short detonation flames)
GENERAL PROPERTIES

Density (g/cc)
Type Density
Granular Dynamite 0.8 to 1.4
Gelatin Dynamite 1.0 to 1.7
Cartridged Water Gel 1.1 to 1.3
Bulk Water Gel 1.1 to 1.6
Air-emplaced ANFO 0.8 to 1.0
Poured ANFO 0.8 to 0.9
Packaged ANFO 0.8 to 1.2
Cartridged Emulsion 1.1 to 1.33
Sensitivity
Type Hazard Performance
Sensitivity Sensitivity
Granular Dynamite Moderate to High Excellent
Gelatin Dynamite Moderate Excellent
Cartridged Water Gel Low Good to Very Good
Bulk Water Gel Low Good to Very Good
Air-emplaced ANFO Low Poor to Good
Poured ANFO Low Poor to Good
Packaged ANFO Low Good to Very Good
Cartridge Emulsion Low Good to Very Good
Packaged Binary Low to Moderate Very Good
Water resistance

Type Resistance
Granular Dynamite Poor to Good
Gelatin Dynamite Good to Excellent
Cartridged Water Gel Very Good
Bulk Water Gel Very Good
Air-emplaced ANFO Poor
Poured ANFO Poor
Packaged ANFO Very Good
(if package is not
broken or torn)
Cartridge Emulsion Very Good
Packaged Binary Poor to Good
Borehole Diameter
Type 2" or less 2" to 4" 4"+
Granulate Dynamite USE N/A N/A
Gelatin Dynamite USE N/A N/A
Cartridge Water Gel USE USE N/A
Bulk Water Gel N/A USE N/A
Air-emplaced ANFO NR USE USE
Poured ANFO NR USE USE
Packaged ANFO N/A USE USE
Cartridged Emulsion USE USE N/A
Packaged Binary USE USE N/A
 Topic 2 –INITIATORS AND INITIATION SYSTEMS
INITIATORS
 Devices with high energy explosive, which, upon receiving
mechanical or electrical impulse, do produce detonation or
deflagration action that starts up the detonation of main explosive.

INITIATION SYSTEMS
 Initiation systems are initiators plus other devices which are used to
start up blasting rounds.
 The systems are: Safety fuse; Normal detonating cord; Low power
detonating cord plus electric detonator; Non electric (NONEL);
Electric blasting system, shock tube system, etc.
EXPLOSIVE INITIATION SYSTEMS
 Explosive Initiation Systems comprise the followings
1. initial energy sources; includes blasting machine-electric
system, shot shell primer -shock tube system, fuse lighter-cap
and fuse.
2. energy distribution network-distributes energy to blastholes:
wire - electric system, tube – shock tube system, fuse - cap and
fuse
3. blasthole initiators relay energy down blasthole
4. booster - unit of cap sensitive explosive that detonates the
main explosive charge
 Some of the explosive initiation systems are:
a. Safety fuse (cap and fuse) system
b. Normal detonating cord system
c. Non electric (NONEL) or ( shock tube) system
d. Electric system
e. Electronic detonators system
f. Low power detonating cord plus electric detonator system, etc.
THE SAFETY FUSE SYSTEM
 It is detonating cord which is loaded with LE (black powder);
 It is a thermal cord which is initiated by fire ignition.
Main Elements
 Safety fuse cord;
 Safety fuse detonator
 Primer (booster or dynamite stick);
 Ignition cord;
 Fuse lighter.
System Assembly
1. Fuse cord is measured to the desired length, cut off squarely with a
sharp knife, inserted to the open end of detonator fuse and the union
crimped;
2. The primer is made by making a hole in the dynamite stick and the
detonator (which is connected with safety fuse) inserted in it;
3. The end of safety fuse with primer is dropped into the blast-hole;
4. Charging of the blast-hole with main charge is completed and once
ready for firing, the other end of safety fuse protruding on the surface is
ignited with a lighter;
5. The blaster walks/drives to the demarcated safe hiding position.
Firing several rounds of holes is carried out as follows:
1. If it is required to blast holes on delayed sequence, holes to be blasted
first are given shorter and the ones to be fired late are given longer safety
fuse lengths;
2. If holes are closely positioned, safety fuse ends protruding on the surface
from individual holes could all be inserted into the one end of ignition
cartridge (containing ignition material at its other end).
3. If boreholes are positioned far-way from one another, the ends of safety
fuses from separate boreholes are attached on an igniter cord with the
help of special metal connectors.
The final stage of preparations for firing a package of explosive charges.
ELECTRIC BLASTING SYSTEM

 Based on the usage of electric detonators with ac or dc power sources

Main Elements
1. Electric detonator:
- It is made of bridge wire, ignition explosive and base explosive;
- Electricity is sent in through copper wires to heat the bridge wire;
- Heated up, the bridge wire starts a chain of explosive burning within the detonator
shell;
- The burning of explosive within detonator shell initiates the high explosive base charge
within the detonator shell;
- The explosion of the high explosive base charge within the detonator shell ignites a
cap sensitive explosive charge.
2. Delay detonator:
- Made by inserting a delay element between igniting charge and the main charge in
the bottom of detonator shell;
- Time of firing the main charge is regulated by varying the length of delay element.
- Short delays are used in surface blasting operations, while longer delays are used
underground where blasting conditions are more confined.
Types of delay detonators:
i. Regular: Delay interval is in seconds
ii. Millisecond: Delay interval varies from 25 to 1000 ms.
System Assembly
Making a primer with electric detonators:

 Hole for detonator is made in the priming stick of dynamite with the pointed handle
of detonator crimper, wooden stick or copper needle;

 Detonator inserted into the hole and tied up.

Designing electric initiation systems

 It is important to know the resistance of the circuit in order to ensure the

sufficiency of firing current in terms of amperage and voltage;

 Electric computations are carried out based on ohm’s law for ac power source.
Types of Electric Shot Firing Circuits

 The methods of connecting electric detonators to the power source are the series,
parallel, series in parallel and parallel in series.

Series connection Parallel connection

 The series circuit
- Entire current passes through each detonator;
- As soon as one fires, the current is cut off;
- In a circuit with more than 50 detonators there may be enough
leakages through ground to make center holes to misfire.
 The parallel circuit
- Is preferable if there is enough power;
- Each detonator is fired independently but greater amperage is
required.
 The series in parallel circuit
- If the voltage of power circuit is not high enough to fire detonators
in a series commutation, the required voltage is reduced by
arranging the detonators in a series-in-parallel, but there should
be the same number of detonators in each series.
 Parallel in series circuit
- Rarely used but reduces the number of detonators in straight
series commutation
NON ELECTRIC SYSTEM - NONEL (SHOCK TUBE SYSTEM)
 Include a cap similar to that of an electric cap but they are
connected to plastic tubing or a transmission line that carries an
initiation (shock and heat) to initiate the cap.
 The energy source in the tubing is either a gas mixture or an internal
coating of special explosive.
 Non-electric tubing is not used in underground coal or gassy mines
as it carries an open flame.
 The plastic tube itself does not detonate; therefore, the only noise
source is the cap itself. (less noisy than typical detonating cord)
 Caps and tubes of varying lengths are connected with special
connectors between holes to configure unique blast pattern arrays.
 tube can be layered for extra durability
 Relatively simple to use
 Cannot be checked for in hole continuity
Generic Design detonators
 ELECTRONIC DETONATORS
 Developed in late 1980’s to present by all major manufacturers
 ED has an electronic counter on a microchip in place of the pyrotechnic delay
charge, and a capacitor to supply the discharge energy for ignition
 Very accurate – Higher timing precision (10 µs instead of 1-10 ms delay
scatter)
 Currently relatively high priced
 Greater safety against accidental ignition (coded firing signal)

Electronic detonator Electronic delay systems

 BLASTING MACHINES
- They are of different sizes (from one which fires 1 detonator to one which
will fire from 50 up to100 detonators;
- They generate high voltage with low amperage and therefore fit for series
connections.
 Circuits Testing can be done by
1. CIRCUIT TESTER (GALVANOMETER):
2. RHEOSTAT
3. OHMMETER – GALVANOMETER

 CIRCUIT TESTER (GALVANOMETER):

- Used to test circuits before firing;
- It is equipped with silver chloride cells generating a very small current such
that a single detonator could be tested;
- In testing a cap or blasting circuit, the terminals of circuit are touched to the
posts of instrument and if there is no deflection of needle, the circuit is
broken;
- Circuit should be tested from safe distance;
- Detonator should be placed inside a pipe for safe testing.
 RHEOSTAT
- Used for testing blasting machine to see that it is in full working order;
- To use the rheostat, lead wires are connected to the posts on the blasting
machine;
- The end of one of these wires is connected to the rheostat at the proper post, the
other wire is connected to the other post of the rheostat but with one blasting cap
in series and another one in parallel;
- If both caps fire when machine is operated, the current generated is double the
amount necessary to fire one cap and thus affords a sufficient factor of safety to
ensure firing the circuit.

 OHMMETER – GALVANOMETER
- Used for measuring resistance of an electric blasting circuit;
- The instrument is connected in the circuit and adjusted until a point of balance is
reached on the galvanometer;
- The resistance of the circuit is then read on the calibrated scale;
- Using it one should be at safe distance from the explosive charge.
-
 Borehole Stemming Materials
- Stemming materials must have the following characteristics:
- High coefficient of friction, heavy and good strength;
- Increasing large fractions in stemming materials densely compacted will raise the
coefficient of friction but also the leakage of the gaseous products of hole charge
explosion.
 Explosives trucks bring in explosives to load the holes
 Same vehicle for explosives and detonator? (may be permitted)
• Blasters are required to have a state issued blasters certification
• Before loading begins all unnecessary personnel and equipment
are removed from the blast site, the site secured and warning
signs posted
 Loading a booster+ detonator + cord down to the hole
 Only workers thoroughly experienced in handling explosives shall be
permitted to supervise, handle, haul, or detonate explosives.
Loading of explosives –Tamping stick
The loaders fill the holes with the amount of explosives that has been
determined –Pneumatic loading
Blaster-in-Charge
Blasting Shelter
 After the holes have been drilled, explosives are loaded in them and they are
joined. Once connected to the blasting maching, the holes can be shot with a
push of a button.
 The blasting machine or the firing key should be securely kept by the blaster
during the entire process of loading and hook up to prevent any unintentional
detonation.
THE THEORIES AND GENERAL PRINCIPLES OF ROCK BLASTING

i. Strength of explosive charge should be proportioned to rock

resistance;
ii. Burden should be proportioned to strength of explosive and the
rock resistance;
iii. Blasting should be designed in a way which should have two or
more free faces for the next shorts to blast on. This enables to
economize explosive consumption;
iv. It is more economical blasting in a system of regular faces and
benches than blasting in irregular manner;
v. Simultaneous firing of several shots grouped closely together
requires less powder factor than if shots are fired singly;
vi. Explosive should not be used more than necessary. Breaking the
rock too small or throwing it further than is needed may be a waste
of explosive.
Topic 3- EXPLOSIVES ENGINEERING
OXYGEN BALANCE AND THE FORMULATION OF EXPLOSIVES
 OB = 100 [OP - OR]/ M %
Where:
 OR – The no. of gram-atoms of oxygen that is required for the complete
oxidation of fuel elements in the explosive;
 OP – The no. of gram atoms of the oxygen content in the explosive;
 M – The molecular weight of explosive in grams.
 Consequently, if the composition of explosive is presented as: CaHbNcOdAle,

 3e 
OB = d   2a    n
b
 100,
  2 2 
m
 Where: n - the atomic weight of oxygen, (n = 16); m - the molecular weight of
explosive.
 The Oxygen Balance (OB) of explosive could be as follows:
- Positive if d > 2a + b/2 + (3/2) e;
- Zero if d = 2a + b/2 + (3/2) e;
- Negative if d < 2a + b/2 + (3/2) e.
 Zero OB generates the maximum quantity of energy and minimum quantity of
noxious gases
 (-) OB generates either the poisonous carbon monoxide (with less heat
generation), or pure carbon in the form of soot, acutely reducing the formation of
gases.
 (+) OB, heat generation drops, since that the poisonous nitrogen oxide is formed,
accompanied with the suppression of heat in the reaction.
EXAMPLE ONE
 Calculate the OB of TNT (C7 H5 (NO2)3) whose molecular weight is 227

Solution
C7H5 (NO2)3 ~ C7H5N3O6
a = 7, b = 5, c = 3, d = 6
OB = 16 x 100 [6 – (2x 7+ 5/2)]/ 227 = -74 %

EXAMPLE TWO
 Calculate the OB of Grammonit 30/70. Grammonit 30/70 consists 30%
AN - ammonium Nitrate (NH4NO3) of molecular weight of 80 and 70%
TNT.
Solution
The oxygen balance of NH4NO3:
NH4 NO3 → N2H4O3
= 16 x 100 [3 – (0 +4/2)/ 80 = + 20 %
The oxygen balance of Grammonit 30/70:
0.3 X (+20) + 0.7 X (- 74) = - 45.8 %
Formulation of Commercial Explosives-ZERO OB
 Commercial explosives are prepared to ensure they have zero OB.
 Cartridge explosives, they are given a small positive OB for the
oxidation of their packaging material.
 For underground works explosives shouldn’t liberate noxious gases
higher than 40 liters of carbon monoxide equivalent from the explosive
reaction of 1 kg of explosive.
 For surface mining their oxygen balance are not so strict.
EXAMPLE THREE
 Establish the formulation of ANFO whose oxygen balance is zero, based
on Ammonium Nitrate (OB = + 20 %) and Diesel fuel DF (OB = -320% )
Solution:
The quantity of weight parts of AN requirement for oxidation of 1 weight
part of diesel fuel equals to:
n = OB Dies / OB AN = 320/ 20 = 16
The content of DF in ANFO is:
X = 100/ [1 + n] = 100/ [1 + 16] = 5.9%
Respectively, content of AN in ANFO is:
100 – X = 100 – 5.9 = 94.1 %
Therefore, the formula of ANFO is: 94.1 % AN + 5.9 % DF
EXAMPLE FOUR
 Establish the formulation of explosive with zero OB, based on the AN and
TNT (C7 H 5 (NO 2) 3. The oxygen balance of TNT is -74 % and the molecular
weight is 227. The oxygen balance of AN explosive is + 20 % and its
molecular weight is 80. Establish also its molecular formula.
Solution:
The composition of the explosive mixture should fulfill condition:
X (-74 %) + (100 – X) 20 % = 0,
Where, X – the content of TNT in the mixture, %.
Solving the above equation indicates that:
X ≈ 21 % And
(100 – X) = 79 %.
 Such composition of the explosive mixture corresponds to grammonit 79/21
and Ammonit standard explosives.
 Let’s express the number AN moles through y, and the number of TNT moles
through z.
Then from the relationship: [80 Y]/ [227 z] = 79/21,
We obtain: y = 79 x 227z / [21 x 80] = 10.7z
and letting z = 1, gives: y = 10.7
Consequently, the molecular formula of Grammonit is:
Z + 10.7 y = C7H5 (NO2)3 + 10.7 NH4NO3
EXAMPLE FIVE
 Establish the molecular formula of Granulit AC-8, which has the
following composition: 89 % AN (NH4NO3); 3% straw oil (C16H34) of
molecular weight is 226; 8 % Aluminium powder (Al) whose
molecular weight is 27.
Solution
 Let’s express the number of straw oil moles through x, AN moles
through y, Aluminium powder moles through z.
 Then, we could write the chemical formula as: Y
yNH4NO3 + x C16H34 + z Al
 In correspondence to weight compositions we could write the
following relationship:
[80 y]/ [226 x] = 89/3; [27 z]/ [226 x] = 8/3.
 From there we obtain:
Y = 83.9 x; z = 22.4 x
 If we express x = 1, the molecular formula of Granulit AC – 8 will be
as follows:
83.9 NH4NO3 + C16H34 + 22.4 Al
 Homeworks
i) Establish the oxygen balance of “TEN” explosive whose chemical formula is
C (CH2O.NO2)4 and molecular weight is 316;

ii) Calculate the oxygen balance of a permissible explosive which is made of 56

% AN; 9 % TNT; 3 % wooden pulp and 32 % NaCl;

iii) Calculate the oxygen balance of a permissible “Ammonit”

explosive which is made of 70 % AN; 18 % TNT and 12 % NaCl;

iv) Establish the chemical formula of Grammonit 50/50, made of 50 % AN and

50 % TNT;

v) Establish the chemical formula of a permissible Ammonit

explosive made of 64 % AN, 16 % TNT and 20 % NaCl;

vi) Establish the percentage contents of Al and TNT required to make their
explosive mixture of zero oxygen balance.
 Topic 4: EXPLOSIVE THERMOCHEMISTRY
 The Heat, volume, temperature and pressure of the gaseous products of
explosion are determined by their composition and quantity.
 The Oxygen Balance of commercial explosives determines the composition of
their products explosion and the conversion reactions of explosive disintegration.
 The reaction structures of the conversion reactions of commercial explosives are
categorized into 3 groups:
GROUP ONE:
 Explosive with oxygen content that is sufficient for the complete oxidation of its
content of fuel elements. Explosives in this group have zero or positive oxygen
balance.
Example: The decomposition reaction of Nitroglycerin:
2C3H5 (ONO2)3 → 6 CO2 + 5 H2O + 3 N 2 + 0.5 O2

GROUP TWO:
 The group includes explosives with oxygen content which is sufficient for the
complete formation of gases.
 The oxygen contained in the explosive, first oxidizes hydrogen into water, carbon
into carbon monoxide, and remaining portion of oxygen combining with CO
forming CO2.

Example: The decomposition reaction of “Ten - explosive”:

C (CH2ONO2)4 → 4 H2O + 3 CO2 + 2 CO + 2 N2
 Hydrogen, methane and other hydrocarbons are also liberated from the explosion
reactions of explosives in this group
GROUP THREE:
 Explosives with a content of oxygen which is not sufficient for complete
formation of gases.
 In this group, Hydrogen is oxidized into water, a portion of carbon is oxidized
into carbon monoxide and free carbon is generated.
Example: The decomposition reaction of TNT:
C7H5 (NO2)3 → 2.5 H2O + 3.5 CO + 3.5 C + 1.5 N2

 More detailed computations have shown that, in the explosions of explosives

in this group, CO2, H2, CH4, NH3 and other combinations are also formed.

THE HEAT OF EXPLOSION

 It is the quantity of heat which is generated from the explosive decomposition
of one mole or one kilogram of explosive.
 The heat of explosive decomposition could be calculated or established
through experiment.
 The standard conditions applied are: temperature 180 C (sometimes 250
C); and pressure 1.01 x 105 Pa.
 Calculation of the heat of explosion is based on the Hess’s Law: heat of
explosion equals to the algebraic sum of the heats of explosive formation and
the heats of its products of explosion.
 Hess triangle
 (1)State of free elements, from which explosive is made; (2) state
corresponds to explosive itself; and the third (3) state corresponds to
explosion products.

1. From free elements, explosive is formed, a reaction characterized by a

positive or a negative heat effect Q1-2;
2. Explosive explodes and heat of explosion Q2-3 is generated;
3. From free elements, explosion products are formed and their heat of
formation Q1-3 generated.
Therefore,
Q1-2 + Q2-3 = Q1-3;
and, the heat of explosion: Q2-3 = Q1-3 - Q 1-2
Tab. 1 The heats of formation of various explosive substances are shown on table below
EXAMPLE SIX
 Calculate the heat of explosion of nitroglycerin, whose explosive
decomposition course is as follows:
C2H5 (ONO2) 3 → 3 CO2 + 2.5 H2O + 1.5 N2 + 0.25 O2
Solution
 From the above table it could be observed as follows:
 The heats of formation at constant volume are:
- For Nitroglycerin: Q1-2 = 351kJ/ mole
- For carbon dioxide (gas): qCO2 = 396 kJ/ mole
- For Water (liquid): qH2O = 241 kJ/ mole
 The heat of formation of the products of explosion:
Q1-3 = QCO2 + QH2O = 3qCO2 + 2.5qH2O
= 3 x 396 + 2.5 x 241 = 1790 kJ/ mole
 The heat of Nitroglycerin explosion could be calculated as follows:
Q2-3 = Q1-3 - Q1-2 = 1790 – 351 = 1439 kJ/mole
 Obtained value (Q2-3 = 1439 kJ/ mole) characterizes the heat of
explosion at constant pressure, Qp
 If the cooling of explosion products goes until it reaches the temperature
of surroundings, 150 C (288K), the heat of explosion at constant volume
Qv is related with Qp as follows:
Qv = Qp + 0.572 n.
 If the cooling of explosion products goes until it reaches the temperature of
surroundings, 25 0 C (298 K), the heat of explosion at constant volume Qv is
related with Qp as follows:
Qv = Qp + 0.592 n.
Where: n = number of moles of the gaseous products of explosion
Let assume the temperature of surroundings is 150 C,
Qv = 1439 + 0.572 (3 + 2.5 + 1.5 + 0.25) = 1439 + 4.2 = 1443 kJ/ mole

The heat generated per unit weight/ kilogram of explosive:

Q1 = 1000 Qv/ me = 1000 x 1443/ 227 = 6357 kJ/kg,
Where: me – the molecular weight of explosive.

THE TEMPERATURE OF EXPLOSION GASES

t = Qv/ Cv, t (0C),
Where, Qv – The heat of explosion , J/mole; Cv – the average heat capacity of
all products of explosion at constant volume, J/ (mole.0C)
Cv = a + bt,
 Where, a, b → coefficients obtained through experiment.

 a  a 2  4Qv b
t=
2b
 Depending on temperature, the heat capacities of certain gases are
determined from the following equations:
 For gases made of two atoms:
Cv = 20.1 + 18.8 x 10-4t J/ (mole.0C)
 For gases made of four atoms:
Cv = 41.9 + 18.8 x 10-4t J/ (mole.0C)
 For water vapour:
Cv = 16.76 + 90 x 10-4t J/ (mole.0C)
 For carbon dioxide:
Cv = 37.7 + 24.3 x 10-4t J/ (mole.0C)
 For solids:
Cv = 26.8 J/ (mole.0C)
 The heat capacity of a gaseous mixture we consider the partial contributions
of each of the components by summing per element to determine the sums
of a, and b in accordance to formula:

t=  a   a   4 b Q.1000
2

2 b
EXAMPLE SEVEN
 Establish the temperature of explosion of nitroglycerin. The heat of
explosion of nitroglycerin is 1443kJ/ mole.
Solution
 Heat capacities of all explosion products (EXAMPLE SIX), based on
the above formulas are:
For 3 CO2: 3 (37.7 + 24.3 x 10-4 t) = 113 + 72.9 x 10-4 t;
For 2.5 H2O: 2.5 (16.76 + 90 x 10-4 t) = 41.9 + 225 x 10-4 t;
Total: CV = 154.9 + 297.9 x 10-4 t;
 Therefore, ∑a = 154.9; ∑b = 297.9 x 10-4
 Using the obtained values on the formula for t, we will obtain:
t = [-154 +√ [23994 + 4 x 297.9 x 10-4 x 1443 x 103]]/ [2 x 297.9 x 10-4]
= 4780, 0C
THE VOLUME OF EXPLOSION GASES
 The volume of explosion gases is established based on the
Avogadro Law which says: “The volume occupied by one gram
molecule of various gases at 00C and pressure 1.01 X 105 Pa is
22.42 X 10-3 m3”.
 The volume of gases (m3/ kg), generated from the explosion 1 kg of
explosive:
Vo = 22.42n1  n2  ...  nk 
m1M 1  m2 M 2  ...  mn M n

Where: n - Quantity of gram molecules of the gaseous products of

explosion;
m - Quantity of gram molecules of the composing parts of
explosive mixture;
M - The molecular weight of the composing parts of explosive
mixture.

EXAMPLE EIGHT
 Calculate the volume of the gaseous products from the explosion of
1 kg of nitroglycerin.
Solution:
The volume of gases:
 For the vaporous state of water,
V0 = 22.42 (3 + 2.5 + 1.5 + 0.25) = 0.616 m3/ kg
1 X 227
 For the liquid state of water,
V0 = 22.42 (3 + 1.5 + 0.25) = 0.469 m3/ kg
(1 X 227)
THE PRESSURE OF EXPLOSION GASES
 The pressure of gases (Pa) resulting from explosion is calculated
based on the Law of Boiler – Marriott and Gay- Lusaka’s as follows:
P= q oVo T ( applicable for ideal gases)
273V
Where: qo - The gaseous atmospheric pressure which equals 1.01 x
105 Pa;
Vo - volume of explosion gases (m3)
T - The temperature of explosion, read from absolute zero, K;
V - The volume of charge chamber, m3.

 For the actual explosive charging densities (0.5 → 1.0 T/ m3), great
role is played by the volume of the individual molecules (co-
volumes) of the products of explosion, which is taken equal to: α =
0.001Vo
P = Po Vo T
273(V- α)
For explosive densities higher than 1g/ cm3: α = 0.0006 Vo
 If the volume of charge chamber V is changed with the loading density
of explosive (Δ =m/V) T/ m3, then for a unit mass (m =1) we will obtain
P = Po Vo T = Po Vo T Δ Pa
273(1/Δ - α) 273(1- α Δ)

EXAMPLE NINE
 Calculate the pressure of explosion gases of Nitroglycerine, whose 1 kg
gives 0.716 m3 of gases at the temperature of 5053 K , for loading
densities of Δ = 0.8 g/ cm3 and Δ = 1.2 X 103 Kg/m3
Solution
 For loading density of 0.8 g/ cm3
P = 1.01 x 105 x 0.716 x 5053 x 0.8 x 103
273 (1 – 0.716 x 0.001 x 0.8 x 103)
= 2.5 x 109 Pa

 For loading density of 1.2 x 103 Kg/ m3

P = 1.01 x 105 x 0.716 x 5053 x 1.2 x 103
273 (1 – 0.716 x 0.0006 x 1.2 x 103)
= 3.3 x 109 Pa
Homework
i) Establish the heat, temperature and volume of the gases of AN explosion,
whose reaction of explosive disintegration is given as:
2NH4NO3 4H2O + 2N2 + O2

ii) Establish the heat, temperature and volume of the gases of nitroglycerin
explosion, whose reaction of explosive disintegration is given as:
C2H4 (ONO2)2 2CO2 + 2H2O + N2

iii) Establish the heat, temperature, the volume of explosion gases and the
pressure of explosion gases of RDX whose borehole loading density is 0.8 T/
m3 and reaction of explosive disintegration given as:
C3H6N6O6 3H2O + 3CO + 3N2

 N.B. Use table in solving the problems.

THE STRENGTH OF EXPLOSIVES AND THE WORK OF EXPLOSION

Theory: The heat generated from the explosion of explosive in a constant

volume without doing external work, represents the total energy of the chemical
reaction;
 The transformation of heat into mechanical work is accompanied with
significant losses;
 The energy of explosive (less chemical losses), generated at the moment of
explosion in the form of heat constitutes the total – actual heat energy;
 That energy is also not totally realized into mechanical work due to heat e.g
thermodynamic losses;
 The work of explosion could be examined as the work of adiabatic expansion
of the products of explosion AT, (kJ/ kg), to a limit, set by the atmospheric
pressure; calculated as:
 V 
k 1

AT = Q 1    
1

  V2  
Where: Q – The potential energy (the total heat energy) of explosion, kJ/
Kg;
V1 and V2 – The initial and end unit volumes, m3/Kg; k = Cp/ Cv - adiabatic
index;
Cp and Cv – heat capacities at constant pressure and volume respectively.
 Replacing the ratio of unit volumes with the ratio of the initial
explosion pressure P1 to pressure P2, when gases have executed
work A, we could write:

 k 1

A=Q 1   P2 

k

  P1  
 

 For explosions conducted in the air (P2 = 105 Pa) and their total
working capability:
 k 1

1   10 
5
AT = Q 
k
Qq
  P1  
 

 The quantity (q = Q – AT) represents is constant heat, remaining in

the products of explosion when their expansion reaches the limit set
by the atmospheric pressure.
 Having that heat, gases are unable to execute work since that, their
pressure equals to the atmospheric pressure (or pressure of the
surroundings);
 For explosion conducted in a media, e.g., with P2 = 107 Pa:
 k 1

1   10 
7
AT = Q 
k

  P1  
 

 Total work becomes less, remnant heat: q = Q – AT goes up

The coefficient of explosion efficiency, ηT:

 The ratio of the total work to the heat of explosion:

AT/ Q = ηT,

 EXAMPLE TEN : Calculate the total work capability (strength)

and the ratio of total work to the heat of explosion of Ammonit
(std) explosive charge of borehole loading density Δ = 900 Kg/m3
and adiabatic index k =1.24 and the following parameters of
explosive conversion: the volume of explosion gases is 0.86 m3/
kg, the heat of explosion is 3930 kJ/ kg and the explosion
temperature is 2600 0K (see tab 2. below)
Tab. 2
Solution: First, determine the pressure of explosion gases for given
charge loading density:
P = PoVoT Δ = 1.01 x 105 x 0.86 x 2600 x 900
273(1/Δ - α) 273(1 – 0.86 x 0.001 x 900)
= 3.36 x 109Pa
 The total working capability (strength) of explosion:

 k 1

1   1.01x10 
5

k

AT = Q   P 
  


 1.241

1   1.01x10 
5
= 3930 
1.24

  3.36 x10 9  
 
= 3402kJ/Kg

 The overall explosion efficiency for the expansion of the explosion

gases to the limit set by the atmospheric pressure: ηT = AT/ Q=
3402/3930 = 0.866
 Homework Exercises:
Use table 2 given below.
i) Calculate the total working capability (strength) and the coefficient of explosion
efficiency of black powder charge of borehole loading density 0.7 t/ m3.

ii) Calculate the total working capability (strength) and the coefficient of explosion
efficiency of nitroglycerin charge of borehole loading density 1.0 t/ m3.

iii) Calculate the total working capability (strength) and the coefficient of
explosion efficiency of TNT charge of borehole loading density 1.0 t/ m3,
blasted in water at a depth of 100m.

iv) Calculate the total working capability (strength) and the coefficient of
explosion efficiency of ammonal charge of borehole loading density 0.9 t/ m3,
fired in a media of compression strength 3 x 107 Pa,

v) Calculate the total working capability (strength) and the coefficient of explosion
efficiency of RDX charge of borehole loading density 1.0 t/ m3.

vi) Calculate the total working capability (strength) and the coefficient of
explosion efficiency of Dynamite - 62 % charge of borehole loading density
1.1 t/ m3, fired in a media of compression strength 1 x 108 Pa,

vii) Calculate the total working capability (strength) and the coefficient of
explosion efficiency of AN charge of borehole loading density 0.9 t/ m3.
 Topic 5: COMPUTATION OF THE DETONATION STATE
PARAMETERS
Theory:
 The equation of explosion products state for solid explosives is given as:
PVn = Constant,
Where: P – The pressure of explosive, Pa; V – The volume explosion
products, m3; n – The polytrophic index of explosion products (dependent on
the initial
density of explosive):
n 1.3 1.6 2.2 2.8 3.0 3.2 3.4
ρe, 0.1 0.25 0.5 0.75 1.0 1.25 1.75
t/m3

 The pressure of shock wave on the C-J plane is calculated from formula:
PD = ρeD2 , Pa,
(n + 1)
 The explosive density on the detonation shock wave:
ρe' = 4 ρe
3
 The masses movement velocity of the explosion products (m/s) on the C-J
plane: D
V = n 1
 The detonation velocity (D) for gases is given as
 
D = 31.5 2 n 2  1 Qv gives too high results when applied on solid
explosives

 DN = DNf QN
Qref
DN, DNf : the detonation velocity of present explosive and reference
explosive respectively, (m/s);
QN, Q Nf: the explosion heat of present explosive and reference explosive
respectively, (kJ/ kg).
 Reference explosive: Ammonit No.6ЖВ or Grammonit 79/20 which have the
following characteristics: Explosion heat: QNf: 4315.7 kJ/ kg; Detonation
velocity DNf: 3600 m/s; Loading density ΔNf: 1.0 t/m3.
 For other borehole charge loading densities, detonation velocity is calculated
from the formula: D = DNf + DNf (ρe – 1)
EXAMPLE ELEVEN
 Calculate the detonation parameters of aquatol 65/35 of charge loading density
1.45 t/m3 and explosion heat Q = 3854.8 kJ/ kg.
Solution
 Let’s calculate the detonation velocity of Ammonite No. 6 reference explosive at
the loading density of 1.45 t/m3 as:
D = DNf + DNf (ρe – 1)
= 3600 + 3600 (1.45 – 1) = 5220m/s

 Let’s calculate the detonation velocity of aquatol:

DN = DNf Q N = 5220 3854.8 = 4933 m/s
4315.7
Qref
 The density of the explosion products:
ρe.p = 4 x ρe = 4/3 x 1.45 = 1.93 g/cm3
3
 The velocity of explosion products movement for:
n = 3.2 + 3.4 – 3.2 x (1.45 -1.25) = 3.28
1.75 – 1.25
D 4933
ω=   1152.6 m/s
n  1 3.28  1
 The detonation pressure on C-J plane:
PD = ρe D2 = 1.45 x 49332 = 8.2 x 106 bars = 8.0 x 105 kgf/ cm2 ≈ 7.8 x 105 Pa
(n + 1) 3.28 +1
Homework
[Use table 2]

i) Establish the detonation parameters of Alumatol charge of loading density1000 kg/ m3 and
explosion heat 5279 kJ/ kg;

ii) Establish the detonation parameters of Grammonal A-45 charge of loading density 900
kg/ m3 and explosion heat 5720 kJ/ kg;

iii) Establish the detonation parameters of ANFO charge of loading

density 800 kg/ m3 and explosion heat 3770 kJ/ kg;

iV) Establish the detonation parameters of Carbonate Э-6 charge of loading density 1100
kg/ m3 and explosion heat 5720 kJ/ kg;

V) Establish the detonation parameters of Ammonium nitrate charge of loading density 900
kg/ m3 and explosion heat 1425 kJ/ kg;
 Topic 6: COMPUTATIONS OF ELECTRIC CHARGE FIRING
CIRCUITS
Theory:
 Electric blasting is based on the following detonator
commutation circuits.
- Series;
- Parallel;
- series-parallel;
- parallel- series.

Series

Parallel series-parallel
 The main condition for the efficient electric blasting is the
provision of non-failure firing of all detonators which are
connected in a network.
 The characteristics of modern detonators are composed of
the following quantities:
i. The resistance of electric detonator:
 Composed of the resistance of detonators bridge
and end wires;
 varies from 2 to 9 Ω (ohm).
i. The maximum non-failure current:
 The upper limit of d.c current, which won’t cause
detonator explosion;
 within limits of 0.18A
i. Ignition impulse (A2.s): kB = I2 tB
 The lowest quantity of current impulse, which
causes detonator explosion:
  Where: I - ignition current, A; tB - the minimum time
of ignition current flow
iv. Sensitivity of electric detonator (A2.S) -1
 the inverse of ignition impulse:
iv. Minimum non-failure current:
 The lowest current quantity, ultimately causing
detonator explosion;
 For direct current: 1.38 A; for alternating current:
1.47 A
iv. Guaranty current quantities for electric detonators.
 For direct current: 1 to 100 detonators: Ig = 1A;
 101 up to 300 detonators: Ig = 1.3A;
 For alternating current: Ig = 2.5 A.
COMPUTATIONS OF ELECTRIC CHARGE FIRING CIRCUITS
 Determination of the quantity of current which will flow through the
bridge filament of the electric detonators.
 Leads to the establishment of the basic assembly scheme of its
elements.
 The condition of non-failure firing of the commutation scheme
requires that guaranty current must flow through all detonators within
the scheme.
 Carried out in accordance to the following sequence:
1. Compute the resistance of mainline, connection and other
wires on the circuit;
2. Calculate the total resistance of the electric charge firing
circuit;
3. Establish the quantity of current available in circuit branches
and in individual detonators
4. Check the condition of non- failure firing.
A. The resistance of main and connection wires, (ohm):
R = 1.1 ρ 4l
πdW2
ρ – Unit length resistance, ohm. mm2/ m;
For copper wire: ρ = 0.0175 ohm. mm2/m;
For aluminum wire: ρ = 0.028 ohm. mm2/m;
l – The length of wire (for mainline wires, length is doubled),
m;
dW – The diameter of wire, mm.
B. The quantity of current in the firing circuit, (A):
I = E/(R + ro)
E - The electromotive force of current power source, V;
R - The total resistance of the firing circuit, ohm;
r0 - The internal resistance of the power source
THE SERIES SCHEME:
 Total resistance in a circuits:
Rs = Rw + nr,
Where: Rw - The resistance of main and connection wires, (ohm);
r - Resistance of individual detonator plus its end wires
(ohm);
n - No. of detonators in the circuit.
 The quantity of current in a single detonator:
i = Is = E/ (Rs + r0)
 Condition of non-failure charge firing:
i ≥ Ig
n
 Rs = Rm + Rc + Ry + nRk + ∑Rd,
1
where: Rm – resistance of main wires, ohm;
Rc – resistance of connection wires, ohm;
Ry – resistance of section wires, ohm;
Rk – resistance of end wires, ohm;
Rd – resistance of a single detonator, ohm.
THE PARALLEL SCHEME:
 Total resistance
Rp = Rw + r/ n
 Amount of current in the circuit:
Ip = E/ (Rp + r0);
 Amount of current in a single detonator:
i = Ip /n
 Condition of non - failure charge firing:
i ≥ Ig
 The required power source voltage:
E ≥ nIg (Rm + Rc + Ry + (Rd + Rk)/n),
THE SERIES IN PARALLEL SCHEME:
 Total resistance:
Rmx = Rm + [Rc + Ry + n1 (Rd + Rk)]/ m,
Where: m – Number of groups which are connected in
parallel;
n – number of detonators which are series
connected in a group.
 Amount of current in the circuit:
Imx = E/ Rmx
 The quantity of current in a single detonator
i = Imx / m
 Condition of non - failure charge firing:
i ≥ Ig
THE PARALLEL IN SERIES SCHEME:
 Total resistance :
Rmx = Rw + m1r/n1 = Rm + Rc + (Ry + Rk + Rd) m1/ n1,
Where: m1 – Number of groups which are connected in series;
n1 – number of detonators which are parallel connected in a
group.
 Amount of current in the circuit:
Imx = E/ Rmx
 The quantity of current in a single detonator
i = Imx / n1
 Condition of non - failure charge firing: i ≥ Ig

N.B. The formulas for the computation of mixed circuits given above
are only applicable on the circuits which have the same no. of
detonators in the groups. Contrary to that, computations become a
bit complicated.
EXAMPLE TWELVE
 Compute the electric charge firing circuit and establish the possibility
of using d.c. power source of electromotive force E = 120 V for the
development of main cross cut of cross section 16 m2. Number of
boreholes is 50. The resistance of detonator with its end wires is 6
ohm. Distance from crosscut face to the power source location is
200m. Also given that:
a. main line wires : are made of double core copper cable, and the
cross section of each wire core is 2.5 mm2;
b. connection wires: Copper wire of cross-section: 1mm2 and
length: 20m (ρ = 0.0175 ohm. mm2/m).
Solution
 Let’s calculate the resistance of mainline wire using formula
RM = ρ 4l x 1.1 = ρ l x 1.1= 0.0175 x 2 x 200 x 1.1 = 3.08 ohm
πdw2 S 2.5
 Resistance of connection wires:
Rc = ρ l x 1.1 = 0.0175 x 20 x 1.1 = 0.385 ohm
S
 Total resistance of the circuit wires:
RT = RM + Rc = 3.08 + 0.385 ≈ 3.465 ≈ 3.5 ohm
For series connection:
 The total resistance of blasting circuit:
Rs = RT + nr = 3.5 + 50 x 6 = 303.5 ohm
 The total current flowing in the circuit:
Is = E/ Rs = 120/303.5 = 0.39 A
 Is < (Ig = 1 A), this means the series commutation does not have
enough current to blast the detonators
For parallel connection:
 Total resistance of blasting circuit
Rp = RT + r/n = 3.5 + 6/50 ≈ 3.6 ohm
 Total current in the circuit:
Ip = E = 120 ≈ 33.5 A
Rp 3.6
 Current available for a single detonator:
i = Ip/n = 33.5/50 = 0.67 A
 Ip < (Ig = 1 A), it implies that parallel commutation also doest have
enough current to blast the detonators.
For parallel in series circuit:
m= 5 series connected groups,
n1 = 10 parallel connected detonators:
 Total resistance of blasting circuit
Rmx = RT + m1r/ n1 = 3.5 + 5x6/10 = 6.5 ohm
 The amount of current in the circuit:
Imx = E/ Rmx = 120/ 6.5 = 18.5A
 In electric detonator:
i = Imx = 18.5 = 1.85A
n1 10
 Since (i = 1.85) > (Ig = 1A), it implies that, the circuit should
work.
 If the No. of detonators is different in branches, current in the
branches is established from the total conductance of all branches.
 Let use four groups of the following no. of detonators per group: 5,
10, 15 and 20.
 The total conductance of all branches:
1= 1 + 1 + 1 + 1 = 1 + 1 + 1 +1 = 25
RB r n1 r n2 r n3 r n4 6x5 6x10 6x15 6x20 360

 Total resistance of all branches:

RB = 360/25 = 14.4 ohm
 Total current in the circuit:
IT = U = 120 = 6.6 A
RT + RB 3.5 + 14.4
 For every unit of conductance (1/360 ohm) will flow the following
amount of current: 6.6/25 = 0.24 A
 The amounts of current in the branches:
I1 = 0.24 x 12 = 2.88 A
I2 = 0.24 x 6 = 1.44 A
I3 = 0.24 x 4 = 0.96 A
I4 = 0.24 x 3 = 0.72 A

 N.B. On the above circuit it is not possible to guarantee the

non – failure explosion of all detonators, since that, explosion
of detonators in the first and second branches could break
the circuit earlier than the detonators in the third and fourth
branches could catch up igniting. Therefore, in mixed circuits
it is important striving to attain uniformity of the resistances of
individual branches.
 Topic 7: SURFACE BLASTING
THE MAIN CONDITIONS OF BLASTING EFFECTIVENESS
 Rock breakage should give smooth working ground floor;
 Degree of rock breakage should be sufficient and economical for the next
processes of rock handling and processing;
 Minimizing sizes increases the productivity of excavators, loaders, transport
system and crushers but also this is associated with higher cost in drilling
and blasting;
 Optimum is that degree of rock breakage (db), defined by size of the average
rock product for which the total cost of rock breakage, handling and
processing is minimum.
 The maximum size of the rock pieces allowed in the blasted rock product is
determined by the parameters of handling facilities - the crushers and other
mechanisms through which the blasted rock product will pass.
For single bucket excavators: lc ≤ 0.8q1/3,
For dump trucks: lc ≤ 0.5Q1/3,
For conveyor transport: lc ≤ (0.5 -0.1) Bc,
For crushers: Ic ≤ 0.75 bc,
Where: q – the volume of excavator bucket, m3;
Q – the volume of dump trucks, m3;
Bc – conveyor width, m;
bc – width of the grizzly holes through which rock product enters the
crusher, m.
 The main disadvantages of excessive oversize
 The costly secondary size reduction;
 The significant reduction of excavation productivity;
 Breakdowns and reduction of the lifespan of transport
equipments.

METHODS OF SURFACE BLASTING

1. Mudcap Boulder Blasting
 For blasting oversize (boulders) and other blasting support
works
 Unconfined charges placed on boulders and subsequently
detonated produce shock energy which will be transmitted
into the boulder at the point of contact between the charge
and the boulder.
 Since most of the charge is not in contact with the boulder,
the majority of the useful explosive energy travels out into
space cause excessive air blast
 Years ago it was found that a thin layer of mud placed on the
boulder with the explosive cartridges pressed into this mud and
subsequently covered by mud causes the explosive charge to exert
more downward force into the boulder than if the mud was not used.
 The mud forms a wave trap, whereby some of the wasted shock
energy, which would normally go off into space, is reflected back into
the boulder
 Hole depth: h = (0.25 → 0.5) hbmad,
 Where: hbmad – boulder thickness, m.
2. Small diameter Hole Short Blasting
 Diameter ≤ 75mm;
 Length ≤ 5m;
 For mining thin valuable ores;
 For breaking boulders;
 For smoothening of bench floors and slopes.
3. Large Diameter hole Short Blasting
 It is the main method for production blasting in surface mining.
 Selection of the diameter of boreholes means selection of the
size of drilling machine and vise versa and is influenced by
- rock crushability,
- the cost of drilling and
- the detonation characteristics of explosive, etc.
Comparison of the effectiveness of small and large diameter
boreholes
 Degree of rock fragmentation is higher for small than in large
diameter holes;
 The cost and productivity of drilling is higher for small than in large
diameter holes.
SURFACE BLASTING EXPLOSIVES
Explosives
 Dry blasting agents : AN based;
 Wet blasting agents: Slurries, emulsions and heavy ANFO.
Selection of the type of explosive;
1. Physical Selection Parameters
- Density- hard massive rock - high density, high VOD explosive
- soft / structured rock - low density, low VOD explosive
- Sensitivity- requires/ doesn't requires booster
- Water resistance
2. Detonation Performance Selection Parameters
- Energy by weight (Absolute Weight Strength - AWS)
- Energy by volume (Absolute bulk strength - ABS)
- Detonation velocity (VOD)
- Detonation pressure
3. Site Specific Selection Parameters
- Rock type - massive rock generally requires products with
higher detonation velocities for optimum
breakage
- fractured rock requires more gas for
displacement
- cracks or voids can require bagged product to
prevent overloading of borehole
- Water – require water resistance explosives
- Explosive cost - compare similar explosives based on $ per
kg and kj of energy provided per $
Rock Blastability on Surface mines
 The difficulty of explosive rock crushing is characterized as the powder
factor of explosive requirement
 Std. powder factor requirement :
qst  0.2Gc  Gm  GT   0.002

Where: γ – rock density (kg/m3);

Gc – upper limit of rock resistance to compression stress (Gc: 0.1→ 450 MPa);
Gm – limit of rock resistance to shear stress (0.1→ 750 MPa);
GT – limit of rock resistance to tensile stress (0→ 45 MPa).
Gc: Gm: GT = 1: 0.3: 0.1
 The above formula is used for comparative purposes and in the
preliminary estimation of materials requirement for the development of a
surface mine
 The actual powder factor requirement for production blasts is influenced
by the following factors:
- strength of explosive;
- the degree of rock jointing;
- charge form;
- bench height;
- the number of free faces
DETERMINATION OF BLAST PARAMETERS - Rules Of Thumb
1. Bench height, H
• Bench heights are normally dictated by site parameters
• If the height (metres) is not predetermined then it should be
greater than the charge diameter (mm) divided by 17 to achieve
good energy distribution
• Drill deviation can be a problem when the bench height is more
than four times the burden dimension

2. Charge diameter/ blasthole diameter , dc

• The selection of the blasthole diameter is a key factor in efficient
blasting
• The maximum suggested charge diameter (mm) should be equal
to the bench height (m) multiplied by 17
• The use of blasthole diameters greater than the suggested
maximum can result in hole deviation and improper energy
distribution
 As blasthole diameters increase the cost of drilling, loading, and
explosives generally decrease
 Smaller blastholes distribute the explosive energy more uniformly
than large blastholes

Acceptable bench height ranges

3. Burden
• Distance from the blasthole to the nearest free face orientated 90
degrees to the rows or blast holes
• Two types of burden which are
1. Face burden, W
2. Burden between rows, b

• Burdens are normally equal (20 40)dc t (for example the burden
for a 165 mm charge would range from 3.3 to 6.6 m)
• Important factors that should be considered during burden selection
include: bench height, rock hardness, structure, explosive used,
desired displacement and the fragmentation required
 Burden dimension should be perpendicular to the desired direction
of displacement
4. Spacing, a
 Distance between blastholes perpendicular to the burden

 Normally ranges from (1 1.8)b

 (industry average ≈ 1.2 x b)
 Optimum energy distribution results when the spacing equals the
burden dimension times 1.15, and the pattern is laid out in a
staggered configuration
5. Subdrilling, ln

 The distance the blasthole is drilled below grade

 Equal to the (3 15)dc with the average length about 7dc

 Dipping or blocky structures (traprock) typically require

more subdrilling (10 15)dc.
 Often horizontally bedded sedimentary structures or soft
seam do not require any subdrill.
 Subdrilling should be minimized as much as possible to
reduce damage below grade and control cost.
6. Stemming l3
 Inert material placed in the blasthole on top of the explosive to confine
energy
 Typically equal to a range of (20 35) dc OR (0.7 → 1.0)W
depending on the charge, stemming type, burden and rockmass
strength.
 If the stemming is less than 18dc then excessive flyrock and premature
venting may occur .
 Wet blastholes require more stemming for confinement than dry
blastholes
 Field trials should be conducted to determine the correct stemming
length when flyrock can not be tolerated

7. Pattern layout
 Square or staggered square
 Rectangular or staggered rectangular
 Slightly rectangular, staggered patterns provide the best explosive
energy distribution (equilateral triangle)
8. Burden stiffness ratio

 If ratio is less than 2 then the rockmass will be stiff, harder to

fracture.
 Low stiffness ratios require relatively higher energy factors to
produce uniform fragmentation
 Low stiffness ratios can produce higher vibration levels and
excessive overbreak
 The stiffness ratio can be improved by using smaller charge
diameters or greater bench heights
The relationship between stiffness ratio and energy distribution
9. Hole charge design- uniform or decked
Decking
 Reducing the charge weight in the blasthole by placing sections of inert
material within the powder column
 Inert material can be made up of drill cuttings or even air as in the case
of air decking
 Minimum decking length required to adequately separate individually
delayed charges is about 14 charge diameters and a primer is required
for each individual charge.
 Air decking can reduce the amount of explosives needed to achieve
good results by efficiently utilizing the available explosive energy
10. Angle drilling consideration
 For inclined borehole charges, rock resistance to blasting is constant along
all bench height, and rock breaks along the line of boreholes, leading to the
following
Advantages
 Better energy distribution
 Improved fragmentation;
 Reduced overbreak
 Better floor control

Drilling 30 degree angled blastholes

Disadvantages
 Requires that close attention be paid to drill set-up
 Generally shorter bit life
 Requires wider drill benches
 Greater hole deviation
 Higher drilling cost per metre
 Requires skilled drillers
 Angled holes in soft material tend to collapse

NB: drill orientation to the free face must be maintained at 90

degrees
BASIC BLAST DESIGN CALCULATION

1. Volume Calculations
 Bank cubic metres (bcms) per hole calculation
= burden x spacing x bench height (B x S x BH)
 Converting bcms to tonnes
= bcms multiplied by the rock density (g/cc)
2. Charging Calculations
 Loading Density (kg of explosive per metre of borehole)
= .000785 x explo. density x (explo.diameter) 2

 The maximum hole charge capacity, Kg:

= charging length x loading density
 Explosive Energy (kcal per kg of explosive) same as explosive
AWS (Absolute Weight Strength) (j/g)
 Loading Energy (kj of energy per metre of hole)
= explosive energy (kcal per kg) x loading density
3. Powder Factor Calculation
 Kilograms of explosive per bank cubic metre
= loading density x explosive column length
bcms per hole
- Good method for tracking costs and relative performance of one
type of explosive if all other factors remain constant
4. Energy Factor Calculation
 Energy of explosive per cubic metre
= loading energy x explosive column length
cubic metres per hole
 Energy of explosive per tonne of rock
= loading energy x explosive column length
tonnes of material per hole
7.1 TUTORIAL COMPUTATIONS IN
SURFACE BLASTING
The Design Parameters of Borehole charges- Tutorial

Parameters
lc - borehole length
lexp - charging length
Hy - bench height
ln - subdrill length
ls - stemming length
W - line of least resistance from the bench toe/ BURDEN
b- burden
a - spacing
c - Distance along the perpendicular line between the upper edge of bench and
the line along boreholes on the first row
The main formulas of blasting rounds charge design in Surface
blasting
1. The weight of rock blasted from one borehole charge : Q = qaWH, kg
Where: q – Powder factor of explosive consumption, kg/m3;
a – spacing, m;
W – Value of the line of least resistance at bench toe, face
burden, m;
H – Bench height, m.
2. Burden at bench toe – W:
W =0.9√(P/mq): inclined or vertical holes (Most common);
W = 2mdc√ (Δe/q): for inclined or vertical holes;
W = 53 krcdc √ Δ/ρ: for inclined holes.
 For vertical holes – Often:
W = √(0.25P2 + 4qoPHyLc) – 0.5P:
2qoHy
Where: krc = coefficient accounting for intensity of rock jointing:
for large size blocked rocks: krc = 1;
for fine, small size blocked rocks: krc = 1.2.
3. W is checked for the fulfillment of condition for the safe positioning of the
drilling machine; Wmin ≤ W ≤ Wmax,
Wmin  H  cot    c  H  cot    3

Δe
Wmax  53k β d c
γ

Where: γ – rock density, kg/m3;

e – powder factor conversion factor (the coefficient of explosive relative
strength);
α – slope inclination, 0;
dc – charge diameter, m;
kβ – coefficient accounting for rock blastability:
For easily blasted rocks: kβ = 1.2;
For rocks of average blast-ability: kβ = 1.1;
For rocks of difficult blastability: kβ = 1.0;
Δ – borehole charging density, kg/m3:
For manual charging: Δ = 900 kg/m3;
For mechanized charging of dry blasting agents: Δ =1000 kg/m 3;
For wet blasting agents: Δ = 1200-1400 kg/m3
4. The Relative Distance between Boreholes – The Coefficient
of Boreholes Closeness on First Row: m = a/ W
m = 1.1 1.2 for rocks of high blastability;
m = 1.0 1.1 for rocks of average blastability;
m = 0.85 1.0 for rocks of difficult blastability.
5. P – Borehole charge loading density, kg/m;
P = d2∆,
4
∆ - explosive density
6. The spacing of Boreholes on a Row

a = mW, m
7. The Spacing between Borehole Rows
b = (0.85 1.0) a
8. The volume of rock blasted from one borehole charge:
V = aWH, m3
9. The volume of rock blasted from 1 m of borehole length:
B = aWH/Lc,
Where: Lc – borehole length, m.
10. Distance along the perpendicular line between the upper edge of
bench and the line along boreholes on the first row:
C = W - H cot (α),
Where: α - The angle of bench slope inclination, degrees.
Usually: C ≥ 3.
11. Borehole charge holding capacity, kg:
QH = [Lc – ls] P,
Where: ls –stemming length, m: (ls) varies from (20 → 35) dc

12. Borehole charging length column:

Lexp = QH/P, m
13. Subdrill length ln = (10 → 15) dc or 0.5qW

14. Borehole length Lc = H +ln

15. The powder factor of explosive requirement:
q = 0.47eγ [de + 0.2)[f]1/4[0.5/dH]2/5/2.6,
Where: de - The average size of natural segregations in the rock in situ, m;
f - The coefficient of rock hardness per scaling Protodyakonov:
The average size of natural segregations in rocks of different jointing
category is given on Table below

 When using explosives other than the standard (reference) explosives:

qN = qe,
Where: q – The powder factor requirement of standard explosive, kg/m3;
e – The coefficient of the relative strength of the new explosive;
qN – The powder factor requirement of the new explosive, kg/m3.
EXAMPLE-THE PARAMETERS OF BLASTING ROUNDS CHARGE
DESIGN

Rocks of density 3 t/m3, hardness (strength) f =16 and of III -rd

category of jointing are being blasted using borehole charges of
diameter 214 mm, bench inclination is 850, bench height is 12 m.
Charging is mechanized (Δ = 1.0). Compute the parameters of blasting
if it is desired to generate oversize +1000mm (d4) equals to zero when
using granulit explosive AC- 4 (e = 0.98). The characteristics of rock
jointing are given on Table below
MONITORING AND OPTIMIZATION OF THE PARAMETERS
OF BLASTING ROUNDS CHARGE DESIGN - TROUBLE
SHOOTING EXERCISES
A. Excessive Generation of Oversize Caused by Irrational Charging of
the Rock In-situ and/ Insufficient Powder Factor
The sequence of assessment and adjustment is as follows:
 Checks on the rationality of charge placement in massif -computations:
Borehole loading density (P): P = d2∆,
4
 Borehole charge capacity (Q,kg): Q = qaWHy
 Charged borehole length (m): LCH= Q/P
 Stemmed borehole length (m):L3 = Hy +Ln -LCH
 Relative distance between holes/coefficient of holes closeness on first
row: m = a/ W
 Conditions of uniform charge placement in massif: m ~1.0
L3 = (20 → 35) dc or
L3 = (1.0→ 0.5)W
 If the above conditions are fulfilled, then charge is uniformly placed in
massif and the main remaining possible cause of excessive oversize is
low powder factor.
B. Optimization of the powder factor
 The percentage of oversize content in massif (V+e) should be
established from photoplanograms of rock massif, taken over
the bench slope face massif just before blasting is carried out
with powder factor (q) and the optimal powder factor
requirement calculated from the formula below:
qo = q V+e/ [V+e – V+H],
Where: V+H – percentage of oversize generation from blast
Carried out with powder factor q.
 Based on the established value of powder factor (qo), the
other parameters of blasting are calculated:
 Borehole length Lc (m): Lc = Hy + 10dc
 Burden at bench toe W (m):
W = √(0.25P2 + 4qoPHyLc) – 0.5P,
2qoHy
W = 0.9√(P/mqo),
EXAMPLES
1. Excessive Generation of Oversize Caused by Insufficient Powder
Factor
Surface mining is blasting a uniform-third category (high jointing) granodiorite-
porphyry rock massif with;
 Coefficient of rock hardness: f = 8 ---10,
 Angle of bench slope inclination: α = 700,
 Bench height: 10 m,
 Diameter of drill bit: 195 mm,
 Sub-grade borehole length: 2 m,
 One row blasting in a grid of: [a x W] ~ [8 x 8] m x m,
 Explosive type: (AN + TNT) 79/21,
 Powder factor: 0.35 kg/m3.
 From experience, oversize generation for that powder factor is V+H 1000 =
25%.
 In order to establish content of oversize in massif, photoplanograms were
taken and as a result, over massif area of S = 15 m2, oversize greater
than1000 mm covered an area of 6.5 m2.
It is required to establish the optimal powder factor to cause reduction of
oversize generation to something not more than 3% and the other parameters of
blast design.
Solution:
Given: α = 700 Hy = 10 m dc = 0.195 m Ln = 2 m a=8m
W = 8 m q = 0.35 kg/m3 V+H 1000 = 25%
 Content of oversize in massif:
V+e1000 = 6.5 x 100 = 43%
15
 Checks on the uniformity of charge placement in massif:
P = πdc2Δ = 22 x 0.195 x 0.195 x 1000 = 30 kg
4 7 4
Q = qaWHy = 0.35 x 8 x 8 x 10 = 224 kg
LCH = Q = 224 = 7.5
P 30
L3 = Hy + Ln - LcH = 10 + 2 – 7.5 = 4.5 m
m = a = 8/8 = 1
W
 Check for uniform charge placement in massif:
If (m =1.0 and 0.5W ≤ L3 ≤1.0W),
m = 1 and 0.5 x 8 ≤ 4.5 ≤ 1 x 8, fulfills.
 The remaining possible cause of the excessive oversize
generated should be insufficient powder factor.
 The optimal powder factor required to overcome oversize
generation to something not more than 3% is calculated from
formula:
qo = q V+e1000/ (V+e1000 - V+H1000)
= 0.35 x 43/ (43 – 25) = 0.83 kg/m3
 Borehole length (m):
Lc = Hy + Ln = Hy + 10dc = 10 + 10 x 0.195 = 12 m
 Burden W (m):
W = (√ (0.25P2 + 4qoPHyLc) – 0.5P)/ (2qoHy)
= (√(0.25 x 302 + 4 x 0.83 x 30 x 10 x 12) – 0.5 x 30)/ (2 x 0.83x 10)
= 6.6 m
 Burden check for fulfillment of drilling safety: W ≥ 3 + HyCotα:
(W = 6.6) ≥ (3 + 10Cot700 = 6.6), implies the condition fulfilled.
2. Excessive Overmilling of Rock Products Caused by Rock
Jointing and Excessive Powder Factor
On a surface mine for flux limestone, a high output of fine fractions (-
100 mm), was observed and constituting a production waste.
 For a powder factor of (AN + TNT) 79/21 explosive mix:
 q1 = 0.5 kg/m3, the average output of fines for the whole mine is
40%
 It is assumed that the high % of fines is caused by irrational
parameters of blasting.
 It is required to establish the degree of over milling (%) and the
rational parameters of blasting to minimize this (%) of fines waste.

Solution:
 Due to the small sizes of rock segregations and fine rock jointing,
measurements of their content on massif were not carried out.
 Instead, a control blast was conducted with a higher powder factor:
q2 = 0.75 kg/m3
 And the output of fines carefully measured, giving: V-H100 = 45%.
 Results from mine record and those obtained from experimental blast
could be presented graphically as shown on figure 9.2:

On figure 9.2 it is observed that:

 The graph of V-e 100 f (q) cuts the y-axis at V-e100 – the % content of fines
<100 mm in massif before blasting,
 V-e100 could be calculated from the following equation:
 (40 – V-e100)/ q1 = (45 – V-e100)/ q2 ~ (40 – V-e100)/0.5 = (45 – V-e100)/0.75,
 Where it is obtained: V-e100 = 30%
 Based on the above results, it could be concluded as follows:
i. In massif (before blasting) there was a content of 30% of over-
milled Products;
ii. The powder factor of 0.5 caused: 40 – 30 = 10% of fines, that is
over-milling of large segregations in massif;
iii. The powder factor of 0.75 caused: 45 – 30 = 15% of fines, that is
over-milling of large segregations in massif.
 Therefore, the main cause of the high generation of fines is the fine
jointing characteristic of massif and nothing we can do to reduce
this.
 The second main cause of fines generation is the high powder factor
and this could be reduced or eliminated through reduction of the
powder factor to its minimum effective value, which will not cause
thresholds on bench floor q3.
 Practical experience in this type of rock has shown that a powder
factor of q3 = 0.3 kg/m3 generates 36% of fines, that is:
(40 – V-e100)/ q1 = (36 – V-e100)/q3,
From which it is established that: q3 = 0.3 kg/m3.
3. Excessive Generation of Oversizes Caused by Irrational
charging of the Rock In-situ
Surface blasting is being carried on granites of the 3rd category of
jointing with oversize content in the rock massif V+e = 100%;
 Blasting is based on vertical holes with one continuous hole
charge:
 Borehole diameter is 110 mm;
 Bench height is 11 m;
 The burden at bench toe W = 6 m;
 Sub-grade borehole length Ln = 1.5 m;
 Holes spacing on first row a = 3 m;
 Stemming length L3 = 1.5 m;
 Oversize content in the blasted product V+H = 40%;
 Thresholds (hard toes) were also observed on bench floor.
 Commutation scheme and charge design are optimal. Without
having to alter the powder factor, find out the solution for
minimizing the oversize generated from blasting
Solution:
Given: V+e = 100% Hy = 11 m dc = 110 mm Ln = 1.5 m
a=3m W = 6 m L3 =1.5 V+H = 40%
 The main cause of the high percentage of oversize
generation is probably caused by the un-uniform placement
of charge in massif since
m = a = 3 = 0.5 << 1
W 6
 Since stemming length (L3 = 1.5 < 0.5 W) is small, it means
that over-size generation is happening along the line of least
resistance.
 Based on the known blasting results and empirical
correlations, we can calculate the radius of rock crushing
perpendicular to charge axis:
 R =( V+e – V+H )W = (100 – 40) x 6 = 3.6 m < W = 6
V+e 100
 Uniform placement of charge in massif is achievable: m = a/
W~1
 Therefore, for optimal charge placement in massif, that is,
when (m = 1), If powder factor is maintained, it will be
necessary to increase the charges spacing (a), and decrease
the burden (W) and crushing Load for a blast-hole will remain
unchanged
Wo = √ Wa = √18 = 4.24 m
 Oversize generation for (W = 4.24) is:
V+H = V+e (1 – R/ Wo) = 100(1 – 3.6/4.2) = 15%,
 This result has been tested and confirmed by blasting tests
with W = 4 m, in which oversize generation was 10% and
thresholds on bench floor also disappeared)
 Therefore to reduce the percentage of oversize without
having to change the powder factor and the quantity of rock
generation from 1 m of hole could be achieved for: burden at
bench toe (W = 4.24 m) and spacing (a = 4.24 m)
4. Excessive Generation of Oversizes Caused by Excessive
Stemming
Surface bench blasting is being conducted on quartzites of 3rd
category jointing.
 Bench height Hy = 10 m;
 Charge grid a x W = 4 x 4 m2;
 Charge diameter dc = 200 mm;
 Sub-grade borehole length Ln = 1.5 m;
 Powder factor q = 0.5 kg/m3;
 Oversize generation V+H = 50 %.
Without having to increase the powder factor, it is required to
establish the cause of large oversize generation and the ways
for its reduction.
Solution:
Given: Hy = 10 m; dc = 200 mm; Ln = 1.5 m; a = 4 m;
W = 4 m; q = 0.5 kg/m3; V+H = 50%.
 Let’s calculate the borehole loading density (P, kg/m) (for hand
charging Δ = 900 kg/m3)
P = πdc2Δ = 22 x 0.2 x 0.2 x 900 = 28 kg/m
4 7 4
 Lets establish the hole charge capacity (QH):
QH = qaWHy = 0.5 x 4 x 4 x 10 = 80 kg
 Let’s establish the length of hole charge column:
LCH = QH/P = 80/28 = 2.8 m
 Let’s establish the stemming length (L3):
L3 = Hy + Ln - LCH = 10 + 1.5 – 2.8 = 8.7 m
L3 = 8.7 > (1.0W = 4)
It means that:
 Cause of the excessive oversize generation is the un-uniform charge
placement in massif;
 Approximately assuming that, most of the oversize is being
generated from the upper parts of the bench, let’s calculate
the radius of crushing along charge axis using formula:
V+H = V+e (1 – (LCB + r) / Hy),
Where: LCB is the charge length above bench toe, m; r = radius
of rock breaking along charge axis:
r = (V+e (Hy - LCB) – V+HHy)/ V+e
 In our case:
Hy = 10 m; LCB = LCH – Ln = 2.8 – 1.5 = 1.3; V+e = 100%,
V+H = 50%
Therefore r = (100(10 – 1.3) – 50 x 10)/100 = 3.7 m
 The radius of crushing perpendicular to charge axis:
R = 2 x r = 2 x 3.7 = 7.4 m.
Lets optimize the parameters;
 Lets calculate the parameters of charge placement, using
approximately formula: W = 0.9√(P/q) = 0.9√(28/0.5) = 7 m
 Let us take (W = a = 7 m), in which (m = a/ W = 1)
 Hole charge capacity will be:
QH = qaWHy = 0.5 x 7 x 7 x 10 = 245 kg
 Charged borehole length LCH becomes:
LCH = QH/P = 245/28 = 8.5 m
 Stemming length L3 becomes:
L3 = Hy + Ln - LCH = 10 + 1.5 – 8.5 = 3 m
 Since (W < R) and (L3 < r), the excessive oversize generation
will be eradicated.
7.2 CHARGE FIRING SCHEMES
CHARGE FIRING SCHEMES

A) Instantaneous multiple row blasting:

 Poor fragmentation
 Wide scattering muck pile.
 Flying rock-debris
 Strong seismic effect lead to ground vibration.

B) Millisecond delay blasting:

 Uniform rock breakage;
 Less faulting of the surrounding rocks in-situ;
 Oversize liberation decreases;
 Explosive consumption is low;
 Scattering of muck pile is less widely.
PURPOSE OF SEQUENTIAL DETONATION (DELAY BLASTING)
1. Control the application of explosive energy during the detonation
of the blast
2. Control explosive energy confinement
3. Maintain the desired explosive energy output

1. Control the application of explosive energy during the detonation of

the blast
- point of initiation is normally where the most relief exists
- point of initiation improve fragmentation and control
displacement and overbreak
2. Control explosive energy confinement
- The blast’s performance will be reduced if the explosive energy
has too little or too much confinement
- Delay times between holes in a row (inter-hole) and between
rows (inter-row) of blastholes are key factors that control the
blast’s dynamic confinement.
- Typically inter-hole delay times range from 1 to 9 ms/m
- Typically inter-row delay times range from 3.0 to 30 ms/m
- Each blast has a specific range of efficient delay intervals that
is dependant on the goal of the blast (fragmentation,
displacement, dilution control, vibration control etc.)
- Too much confinement (short delay intervals) causes the
explosive to find undesirable paths of relief, resulting in
overbreak, poor fragmentation and excessive vibration
- Too little confinement (long delay intervals) permits the
explosive energy to vent, wasting money and potentially
causing flyrock and misfires
3. Maintain the desired explosive energy output
- The energy level must be sufficient to overcome the structural
strength of the rock and permit displacement
- Improper delay intervals can cause low energy output of some
explosives or premature venting
Initial Timing Design Development

1. Define the goal of the blast (ie. vibration control, good

fragmentation, etc.)
2. Prioritize site parameters that affect blast design
- evaluate free face
- determine vibration limitations
- geology - are weak layers such a mud seams present in
the free face, bedding planes and slips.
3. Determine the point of initiation
- point of initiation is normally where the most relief exists
- point of initiation will help control the direction of
displacement
4. Determine the desired direction and amount of
displacement
- the relationship between the inter-hole and inter-row
delay intervals influences the direction of displacement
- Fragmentation can often be enhanced by blasting across the
strike of the major jointing

With the dominant joint orientation shown above which direction

of displacement would provide the best fragmentation? ( 1 2
3 )
 - Potential ore body dilution should also be considered when
determining the direction of displacement

Show the desired direction of displacement require to minimize ore dilution

5. Select the inter-row delay interval based on the site

restrictions and performance goals
6. Select the inter-hole delay interval based on fragmentation or
displacement goal
7. Remember that the timing configurations cannot
compensate for poor blast design
BLAST TIMING AND DELAY CONFIGARATION - Design
Considerations

1. Fragmentation
 Uniform fragmentation generally requires the production
of new free faces during the detonation process .
 Optimum fragmentation in blocky and massive rock
generally occurs when one hole is detonated per delay
and the inter-hole delay is < 1 ms per m of spacing
 The delay between rows should be at least 2 to 3 times
the delay between holes in a row (<2.0 ms/m of burden)
 In highly jointed or highly bedded rock the delay interval
plays a lesser role in the fragmentation of the rockmass.
2. Vibration Control
 Fast inter-hole delay intervals and inter-row delay can
increase ground vibration
 Based on the findings of the study, many regulations have
adopted the “8 ms rule” which states that if charges are not
separated by at least 8 ms their charge weight has to be
added together to estimate potential vibrations

What is the maximum number of holes that fire with less than 8 ms of
delay separation?
What is the maximum number of holes that fire with less than 20 ms of
separation?
3. Muckpile displacement
 The direction of displacement depends path of
least resistance for the explosive energy to follow
 With the proper blast design the delay sequence
can control the direction and degree of
displacement
 Typically longer delay intervals (>12 ms/m of
burden) are required between rows to maximize
displacement
 The type of excavator will often determine the
degree of displacement required which will
dictate the delay interval between rows of
blastholes
4. Wall control
 Too short of delay intervals between holes in a row and between
rows can cause excessive overbreak
 If the delay between blastholes in the back row is less than 6 ms/m
the charges can act together to damage the back wall
5. Geology

 The hardness and structure of the rockmass should be taken

into consideration when developing delay configurations
 Inter-row delay intervals for massive rock typically range from
6 to 30 ms/m
 Inter-row delay intervals for blocky rock typically range from 6
to 18 ms/m
 Inter-row delay intervals for adversely dipping rock typically range from 6 to 9
ms/m
 Soft rock masses can require more time for displacement so more time
should be allowed between rows to overbreak (>4 ms/ft) with the exception
of very soft gypsum or conglomerates which require relatively fast times (<
1.5 ms/ft of burden) to optimize fragmentation
TYPES OF PATTERN CONFIGARATION
takes
advantage
of two free
faces, typically
provides the
best relief
used when no
relief face exists
or to stack the
muckpile in the
center of the
blast
Guidelines for row-to-row initiation are as follows:

 Short delay times cause higher rock piles closer to the face.
 Short delay times cause more backbreak.
 Short delay times cause more violence, air blast and ground
vibration.
 Short delay times have more potential for flyrock.
 Long delay times decrease levels of ground vibration.
 Long delay times decrease the amount of backbreak.
Topic 8: UNDERGROUND BLASTING
Introduction
 Drilling and blasting in the underground mining is carried out
in the
- Mining development e.g. drifts, crosscuts, shafts, raises
and winzes
- Production mining in blocks
 The factors determining the effectiveness of underground
blasting are:
a) Strength of explosive:
- volume of gases produced,
- velocity of detonation,
- the temperature to which the gases are raised at the
time of detonation.
b) Physical properties of explosives
- Water resistance
- Density
- Fume class
- Temperature effect
 c) Physical and mechanical characteristics of rock:
- Strength of the rock
- planes of cleavage and fracture
- the size of excavation

 Those above factors should be considered in

determining
- the number of boreholes in a blast round,
- the direction in which holes are pointed,
- the explosive used,
- manner in which holes are fired,
- and the manner in which rocks break.
BLASTING IN THE DEVELOPMENT OF HEADINGS/TUNNELS
 The General Principle for Underground Blasting is that
- Blasting must carried in a manner that when the first holes
are fired, they break out the rock in a manner which will
create more than one free face for the others;

BLASTING IN DRIFTS AND CROSSCUTS

 Groups of holes are
a) Cut holes:
- The first group of holes to be fired in a drift or cross cut
face to provide new free faces so that the remainder of
holes will break easily.
- Cut holes are normally drilled 0.2 to 0.3 m deeper than
the other holes and charged with a higher amount of
explosive.
b) Reliever Holes:
- These holes helps rock breakage by cut holes and
detonate just after the cut holes;
c) Trim Holes:
- These are the back holes – on the roof of a round, the rib holes –
along the sides of a round and the lifter holes – at the bottom of a
round.
- Trim holes are the last fired system of holes in a blast round,
meant for creation of the predetermined contour of development
d) Breast or Enlarger Holes:
- The holes between the trim and reliever holes,
- They are uniformly distributed on the free face of development
between reliever and trim holes
The main types of cut holes are;
1. Straight Cuts-
i) Coromat Cut
ii) Straight Cut with Two Large Diameter Holes
iii) Straight Cut with One Central Burn Hole
iv) Straight cut with one uncharged central hole of
diameter 45 mm and arrangement of support
holes
2. Angle Cuts -
i) Wedge Cut
ii) Fan Cut (Draw Cut, Toe Cut)
iii) Three hole pyramid
1. Coromat Cut

- Holes are drilled parallel to one another

- One large hole is drilled as a burnhole .
- The number of support holes around the
central cut is determined by the cross section
area of the heading development.
2.Straight Cut with Two Large Diameter Holes

 The most effective of the straight cuts – with two central

holes of diameter 76 mm + support holes,
 Charge amount for support holes is larger because they
operate in conditions of high compression.
3.Straight cut with one uncharged central hole of
diameter 45 mm and arrangement of support holes
1. Wedge Cut (V-cut)

 Blasthole are drilled at an angle to the face (appx. 600) in a

uniform wedge formation so that the axis of symmetry is at
the centre line of the face
 It is the most common of the angle cuts
DESIGNING OF BLASTING ROUNDS FOR HEADINGS DEVELOPMENT

Main parameters of drilling and blasting in underground

blasting :
(i) Borehole diameter,
(ii) Hole depth;
(iii) Type of drilling machine;
(iv) Type of explosive;
(v) Powder factor;
(vi) Explosive consumption for blast round;
(vii) Number of blast holes for round;
(viii) The arrangement of holes on development face;
(ix) Method of initiation;
(x) The consumption of blasting materials per round.
Borehole diameter
- Diameter varies from 36 to 46 mm depend on diameter of
standard commercial cartridges ranging from 28 to 36 mm.
Borehole Depth
- Optimal within the range of 2.5 to 3.5 m
- Established in accordance to the area of heading cross
section and rock hardness as per below table

Heading cross section Hole depth, m, for:

S , m2 f ≤ 12 f > 12
1.5 - 3 2-3 2.5 – 3.5
4.0 - 6 1.5 - 2 2.2 – 2.5
7.0 - 25 1.2 – 1.8 1.5 – 2.2

Type of the Drill Rig

- depends on physical-mechanical properties of rock and the
required drill productivity for predetermined volume of drilling
work.
Type of Explosive
- depends on physical-mechanical properties of rock and
mining conditions of organization
Powder Factor
- determined by the physical-mechanical properties of rock,
- cross section area of development face,
- depth and diameter of boreholes
- type of explosive.
- Usually the powder factor requirement for underground blasting is
established from experience.

Quantity of Explosive per Round

Q = qsl = qV (Kg), kg
Where s - cross section area of development face, l- borehole depth
and q- powder factor.

Number of Boreholes per Round

 The number of holes per round: N = Q/Q1,
Where: Q - quantity of explosive consumed per round; Q1 -
quantity of explosive consumed per hole:
Q1 = d2 LHkρexp,
4
Where: LH - borehole length; d2/4 - cross section area of the explosive
cartridge/ borehole (m2); ρexp - density of explosive cartridge (kg/m3);
k - coefficient of hole filling (k = 0.75).

Boreholes Arrangement on the Heading Face

- As illustrated on previous slides
- Cut holes- drilled 0.2 to 0.3 m deeper than the breaker holes, as a
result are charged with large amounts of explosive
- Breaker holes- drilled and charged in accordance to the parameters
obtained from blasting computations (borehole length and quantity of
charge per hole).
- Trim Holes
- Charges of trim holes are 20% less than the computed quantity of
hole charge. The holes are drilled at a distance of 0.6 – 0.8 m from
one another.
The scheme of boreholes arrangement on the heading face:
1-4 (cut holes); 5-9 (breaker holes); 10-21 (Trim holes)
Tutorial Exercise:
Establish the blasting chart for the development of two way
main cross cut of cross-section area 13.5 m2 (3 x 4.5 m) in
rocks of hardness f = 12.
Solution: Assumptions
- standard explosive AN in cartridges of diameter 30 mm.
- Hole depth – 2 m.
- Work is organized in one round per shift.
- Drilling rate is 110 mm/min.
i. Powder factor requirement for given rock hardness and
cross-section area of crosscut. In accordance to the rock
hardness and the cross section area of the heading: q = 1.7
kg/m3
ii. Total explosive consumption per round: Q = qSLH = 1.7 x
13.5 x 2 = 46 kg.
iii. No. of holes per round: N = Q/Q1:
- The per hole explosive consumption: Q1 = [d2/ 4] LHkρexp;
= 3.14 x 0.03 x0.03 x 2 x 0.75 x1000/4 = 1.07 kg
N = 46/1.07 = 43.
iv. The arrangement of holes on development face
- Lets select a wedge cut of 6 holes. Trim holes are positioned on
the cross cut contour at intervals of 0.8 m and their number is
18.
- The number of reliever + breast holes is 19.
- Computation of the exact consumption of explosive per round:
- Cut holes: Qc = Q1 x 1.2 x Nc = 1.07 x 1.2 x6 = 7.7 kg;
- Breaker holes = (Reliever + breast) holes: QrB = Q1 x NrB = 1.07 x
19 = 20.3 kg;
- Trim holes: Qt = Q1 x 0.8 x Nt = 1.07 x 0.8 x 18 = 15.4 kg;
- Total consumption of explosive: Q = Qc + QrB + Qt = 7.7 + 20.3
+15.4 = 43.4 kg.
 Q = 46 kg differs from Qt = 43.4 kg . The difference could be
eliminated by modifying the charges of the breaker holes
v. Selection of blasting method: Safety fuse
vi. Requirement of detonators: 43 units.
vii. Requirement of safety fuse cord per round: 43 x 3 = 129 m.
viii.Volume of drilling work per shift:
- Cut holes: 6 x 2 x 1.2 = 14.4 m;
- Breaker holes: 19 x 2 x 1.0 = 38 m;
- Trim holes: 18 x 2 x 0.8 = 28.8 m;
- Total: 14.4 + 38 +28.8 = 81.2 m.
ix. Face advancement per round/shift (for coefficient of blasting
efficiency of  = 0.9): l x  = 2.0 x 0.9 = 1.8 m.
x. Volume of blasted rock per shift: V = SLH = 13.5 x 2.0 x 0.9
= 24.3 m3
xi. Explosive consumption per 1 m3 of blasted rock: Q/V =
43.4/24.3 = 1.78 kg/m3:
xii. Drilling consumption per 1 m3 of blasted rock: L /V =
81.2/24.3 = 3.34 m/m3
PRODUCTION BLASTING IN UNDERGROUND

 Metal-bearing orebodies are extracted in underground mines

by various methods, depending on the size, orientation, depth
and geological characteristics of the deposit.
 However, the excavation work is usually divided into two
broad categories; development and production.
 The production work can be subdivided into two categories:
i) Short-hole blasting
ii) Long-hole blasting.
Short-hole blasting
 The diameter and length of shot-holes are usually limited to
43 mm and 4 m respectively.
 Short-hole blasting is usually used in breast stoping for
narrow, tubular orebodies such as gold or platinum reefs.
Example of blast pattern in a South African narrow reef
 About 1.0 to 1.2 m stope width
 Staggered rows of 35 mm diameter blastholes
 The length of the blastholes is about 1.2 m
 The spacing of each hole being 0.5 to 0.6 m.
 The most common types of explosive used include a cartridged nitroglycerine-
based semi-gelatine or emulsion explosive with a composition density of 1.25
g/cm3and a column of pneumatically loaded ANFO with a density of between 0.8
and 0.9 g cm3
 In an average stope, there are about 120 shotholes per panel
 To prevent the hanging wall and footwall being damaged, the present practice is
not to detonate two or more shotholes simultaneously.
 The initiating system is Capped safety fuse and igniter cord
Long-hole blasting
 Basically there are three long hole blasting systems:
i) Ring blasting,
ii) Bench blasting
iii) Vertical crater retreat (VCR).
Ring blasting
 Ring blasting has wide application in massive bodies with
their high rate of extraction at low unit costs.
 The method requires three distinct operations
1. The formation of a tunnel, called the ring drive, from a
sublevel along the axis of the proposed excavation.
2. The excavation of an empty space, called the slot, at the
end of the ring drive, to the full width of the excavation.
3. The drilling of sets of radial holes, called Rings, parallel to
the slot at appropriate spacing and burden.(Normally, the
spacing/burden ratio is about 1.3, but it can be as high as
1.5)
Operations in ring blasting
 It should be noted that the degree of success in ring blasting
depends on the degree of accuracy in designing and drilling
the blasting holes.
 The stemming length should not be more than two-thirds of
blasthole length.
 Also blastholes need to have a variable stemming length in
order to avoid serious overcharging in the ore body close to
the ring drive.
 The blasthole in each ring could be drilled upward as well as
downward
Bench blasting
 Bench blasting is essentially similar to surface excavation
 A development heading is first excavated at the top sublevel to
provide drilling space.
 Then depending on thickness of ore-body and/or availability of
drilling machinery, either vertical or horizontal blastholes are drilled
to increase the height of the excavation

Vertical blastholes
 The blastholes can be from 32 mm up to 250 mm in diameter,
depending on factors such as quality of rock, fragmentation
requirement, etc.

Horizontal blastholes
Vertical Crater Retreat (VCR)
 Vertical or sub-vertical blastholes are drilled downward from the top
level to the bottom level
 A cuboid of ore-body can be excavated from the lower level upward
by a number of horizontal slices using the same blastholes
VCR loaded explosive column
 Topic 9: CONTROLLED BLASTING
Introduction

 Controlled blasting is based on reduced explosive quantities loaded

into holes which are generally smaller in diameter and spaced closer
than the main blast.
 The method is used to control ore break, reduce fractures within
remaining rock walls and reduce ground vibrations.
WALL CONTROL TECHNIQUES
1. Line Drilling
 Closely spaced, unloaded holes which form a natural excavation
line beyond which no rock is to be blasted .
 Borehole diameter ranges from 51-76 mm and spacing from 0.1-
0.3 m.
2. Presplitting - Preshearing
 Based on a line of closely spaced loaded boreholes of diameter
51-101.6 mm and spacing 0.61-1.22 m and fired before the
main charge.
 Charge length is usually 2/3 of borehole length;
 They are drilled along the periphery of an excavation and
initiated before main blast is detonated
 Charges range from 0.06-0.42 kg/m;
 Charge length could be reduced further by 55 % of the
borehole length when blasting in rocks which are highly
fractured.
 The presplit holes should not be stemmed unless airblast
levels must be restricted
 The objectives of line drilling and pres-plitting is to generate
a line of cracks connecting the holes with the intension to
- Terminate the growth of the radial cracks
- Act as a barrier to the shock wave
- Provide an escape route for the explosive gases
 Line drilling produces one of the best final surface - a
smooth, clean face with no backbreak or crest fracture.
However because of its high drilling cost, the method has not
been commonly used in open pit work.
3. Cushion or buffer blasting
 Typically the loading and pattern of the first one or two
rows next to the final wall is reduced
 Charge weight in the toe row is reduced by about 45%
 The toe row’s burden and spacing are typically reduced by
25%
 Only minimal stem is used in the toe row so the explosive
gases can escape vertically
 Usually the toe row adjacent to the final wall is shot last
which will not prevent damage from adjacent rows of
blastholes
 Buffer blast designs normally involve no more than four
rows
 The buffer row should be offset from the presplit row by 40
to 70% the normal burden dimension depending on the
strength and structure of the rockmass.
3. Contour – Trim Blasting
 The method is mainly used in underground blasting
 It is based on explosives of low energy concentration and
characteristics which are favorable for the provision of
smooth explosion action.
 Contour charge is fired after the main charge.
 The distance between contour holes is less than the distance
between other holes in the same rock type.
 Contour blasting is applied in the development of headings,
underground reservoirs, road excavations, etc.
 The advantages of contour blasting are:
- Smooth wall cuts;
- Minimal cracking in the rock in situ.
CONTROL GROUND VIBRATION
 Components of Ground Vibrations
1. Amplitude –
 Can represent velocity (V), acceleration (A) or
displacement (D) depending on the recording
equipment used – typically represents velocity for
blasting measurement
2. Velocity (V) –
 The speed the particles are moving back and forth
 The maximum rate that the particles are moving is
known as peak particle velocity (PPV) and is used
in determining the possibility of damage
 Maximum particle velocity is usually recorded in
millimeters per second (mm/s)
 Over-confined charges can increase particle
velocity
3.Frequency (f)
 The number of times a particle moves back and forth in one
second
 Motion back and forth is called an oscillation or cycle
 The number of oscillations a particle undergoes when
subjected to a vibration wave - measured in cycles per
second or Hertz (Hz)
The relationships between velocity, frequency, acceleration
and displacement are:
Monitoring and control of ground vibration

1. Peak particle velocity prediction formula

where:V = ground vibration as peak particle velocity (mm/s)

K = constant related to the site and rock properties
R = distance between the charge and the point of concern
(m)
Q = maximum instantaneous charge weight (kg)
B = constant related to the site and rock properties
(usually -1.6)
 The K factor will typically vary with confinement as
follows:
- Under confined, K = 500
- Normal confinement, K = 1140
- Over confined, K = 5000
Vibration damage potential
2. Vibration control with the use of a seismic monitor
(Seismographs)
 Assessment of seismic vibrations is based on data obtained
from seismic measurements instruments.
 Data from the seismic monitor can provide information as to
the efficiency of the blast.
 Seismic monitors are the best defense against possible
damage claims.
 There are quite range of Seismographs such as
a. Mechanical Instruments:
- Vibrograph,
- Combigraph
- Ampligraph etc.
 Electrical Instruments:
- Ultra Violet Recorders,
- Vibrocorders, etc.
Seismographs

 Employed in the field to record the levels of blast-induced ground

vibration and airblast.
 Placement and coupling of the vibration sensor are the two most
important factors to ensure accurate ground vibration recordings.
 The sensor should be placed on or in the ground on the side of the
structure towards the blast.
 Where access to the structure and/or property is not available, the
sensor should be placed closer to the blast in undisturbed soil.
CONTROLLING AIR BLAST

 Air blast is air overpressure

 Also full waveform recording instruments (seismograph) can
help determine the source of the air blast in relation to the first
charge detonation
 Air blast control can be done by the use of microphones in
the seismographs.
 Placement of the microphone relative to the structure is the
most important factor.
 The microphone should be placed along the side of the
structure nearest the blast.
FLYROCK CONTROL
 Flyrock is the undesirable throw of rock or debris from a blast
area and is a leading cause of fatalities and equipment
damage from blasting.
Typical Causes
 Blasthole Overloading

Causes:
- Carelessness
- Improper stemming length
- Weak geology, etc.
 Inadequate Face Burden

Causes
a. improper blast design
b. over excavating the face
c. weak face caused by overbreak from
previous blast, etc.
 Inadequate Timing Between Rows Of
Blastholes

Causes
a. poor timing design
b. improper design implementation
c. inaccurate delays
 Excessive Powder Factor

Causes
a. improper design and or implementation
b. too large of charge diameter
c. improper charge density
d. improper stem length
 Secondary Blasting

Causes
a. lack of energy confinement
b. overloaded
  Adverse Geology

Causes
a. weak geological structures( mud seams and joints) poorly confine
explosive energy
b. voids can cause the hole to be overloaded
c. soft overburdens
Solution of the above causes
1. Good shot design is the primary method for
avoiding flyrock
2. But good design cannot completely eliminate
flyrock due to geological inconsistencies so
expect the worst and plan accordingly.
3. Even with proper precautions, flyrock can still
occur
4. Unusually long throw distances should be
recorded and used for determining the proper
blast area to be cleared.
5. All observers of the blast must have adequate
protection from flyrock.
 Part two: DRILLING

Topic 1-DRILLING SYSTEMS

Introduction
 Rock drilling can be defined as rock penetration .
 Rock penetration is method of forming a directional hole in a
rock.
 Rock drills are classified in accordance to their purpose,
mode of attacking the rock and source of power used.
 In accordance to their purpose: Are classified as exploration,
production/blasthole drilling , drainage and degasification,
shaft sinking etc. rock drills.
 In accordance to their mode of attacking the rock leading to
penetration : Are classified as mechanical, thermal ,
chemical , physical and combine rock drills.
 In accordance to their power source used: Are classified as
pneumatic, hydraulic, electrical and combine rock drills.
 Most rock drills are classified based on mode of rock attack
and source of power they employ .
 The mechanical attack category cover by far the majority
(probably 98%) of rock drills used today.
Drilling system
 There are 4 main functional components of a mechanical drilling
system which are
1. Drill rig (energy source)
2. Drill rod/steel (energy transmitter)
3. Bit( energy applicator)
4. Circulation fluid
 These components are related to the utilization of energy by
drilling system in attacking rock in following ways:
a. The drill rig is the prime mover, converting energy in its
original form (fluid, pneumatic, electrical, etc.) into mechanical
energy to actuate the system.
b. The rod transmits energy from the prime mover to the bit
c. The bit is the applier of energy in the system, attacking the
rock mechanically to achieve penetration
d. The fluid cleans the hole, control dusts, cools the bit and at
times stabilizes the hole.
Blasthole drilling
 When drilling a blast hole, a drilling system must perform two
separate function in order to achieve advance into rock
1. Fracture and break
material from the
solid (crushing)
2. Eject the debris formed.
 Good drilling practices include
- carefully monitoring drill-rig operation
parameters,
- taking careful notes on the changes
of geology during drilling
- effective communicating with blasting
crew any unusual Conditions
encountered during drilling that may
affect blasting results or require
changes in hole loading practices.
Mechanical drilling
 The application of mechanical energy to rock can be performed by
either impact (percussion) or simple rotation or both.
 Not all the energy is expended in breaking the rock, some of it is
lost.
 Sources of energy loss include:
- Friction at coupling and other contact points,
- Bit wear
- Noise
- Vibration.
- Flushing (positive work)
- Rotation of steel (positive work)

Classification of mechanical rock drills:

 Rotary (R)
 Rotary-Percussive (R-P)
 Percussive- Rotary (P-R)
 Percussive-turn (P)
Rotary (R) drills
 In rotary drilling, the bit attacks the rock with energy supplied to it by a
rotating drill stem.
 Their process is under the simultaneous action of feed force on the drill
bit which is rotating on hole face

Fc

Mrm

The action of forces and the form of hole face in Rotary Drilling

 Drive: electrical, pneumatic or hydraulic.

The operation principles:
 Characterized by the continuous forward advancement of the drill bit
towards hole face, cutting from it a spiral slice of rock of thickness (h)
under the action of pull down pressure /feed force (Fc) and rotation
torque (Mrm).
Hole cleaning- removal of drill products
 Via the spiral grooves on drilling auger steel
 Compressed air and water.

Areas of favorable application

 In the rocks of minimal abrasiveness and hardness (f = 6 -- 8)
and in rocks of (f > 10) when using drill bits with diamond
armed inserts.

Advantages of rotary drilling.

 High productivity because of the continuity of rock crushing
action.
 Low dust liberation and energy consumption due to the rock
crushing process taking place in thick bites.
Disadvantage of rotary drill
 Limited areas of application. Applied in very soft rocks e.g.
coal and limestone.
Rotary – Percussive drills (R-P) /Top hammer

Fc + Fu

Mrm

The action of forces and form of hole face in R-P drilling

Operation principles
 R-P drilling is composed of percussion impulse (Fu), rotation
torque (Mrm) and significant feed force (Fc),
 The piston drill is fixed on top of the drilling steels and the
rotation mechanism.
Areas of favorable application
1. In rocks of hardness (f = 6 ----14)
2. Only suitable for shallow-small diameter drill holes.

Bit cooling and hole cleaning :

 Accomplished from compressed air or air-water mixture
supplied in through the drill, the central hole in the drill
steel and special holes in the drill bit to cool the drill bit and
flush out rock drill products via the spacing between drill
steel and the walls of drilled hole.
Percussive-Rotary drills (P-R) - DTH

Fu+Fc

Mrm

The action of forces and form of hole face in P-R drilling

Operation principles
 P-R is composed of significant percussion impulse (Fu),
continuous rotation torque (Mrm) and continuous feed force
(Fc)
 The rotation mechanism is fixed on the top of hole surface-
drilling steel;
 The piston drill is fixed on the bottom end of drilling steels and
in contact with the rock on hole face
Bit cooling and hole cleaning :
Accomplished from compressed air or air-water mixture
supplied in through the drill.
Disadvantage of P-R drills
 Since it is difficult to manufacture P-R (DTH) drills of smaller
diameters, these drills are only for deep holes of larger
diameters > 90 mm.
Advantages of P-R/DTH drills over R-P/Top Hammer
 Efficiency in hard rocks types
 With the hammer in the hole drilling vibrations are reduced
 Tend to drill straighter holes at greater depths as compared to
R-P drills
 With the increase of hole depth and the consequent of drilling
steel extension, percussion efficiency is retained in P-R drills
and decreased in R-P drills;
Schematic drawing of three types of drilling
A- Top Hammer (R-P) ; B- DTH (P-R) C- Rotary
a - tip, b- bit, c- rod, d- sleeve, e- drill pipe, f- piston, g- cylinder,
h- percussion mechanism,
i- rotation mechanism, j- flushing.
Percussive – turn rock drills (P)
 The P is based on indentation- that is, the drill bit is placed in contact
with the rock and the periodic- short lived blows of percussion stress
(Fu) applied by the drilling machine on the drill bit, causing its inserts
or buttons to penetrate the rock through a thickness (h)

Mrm
Fu +Fc

The action of forces and the form of hole face in Percussive-turn rock drilling

The operation principles:

 The feed force or thrust due to the weight of the drill (Fc) is very
small and only sufficient to maintain the drill bit - rock contact during
the moments of piston blow.
 The rotation torque (Mrm) is also very small, only sufficient to turn
drill bit through the required angle to enable bit inserts or buttons
operate on fresh position of hole face after every consecutive blow.
 The rotation mechanism turns round the drill steel and the drill bit
through angle β = 10 200 for insert/brazed bits or β = 5 70 for
button bits, enabling them to work on new position of the hole face in
the following piston blow.
Bit cooling and hole cleaning
 Accomplished from compressed air or air-water mixture supplied in
through the drill.

Main advantage of Percussive – Drills

 Capable of drilling the hardest rocks (f = 8 20)

Disadvantages of Percussive-Turn drills

 Low productivity: More time spent on drill bit movement towards hole
face, retreat from it and turn around through the angle β than time
spent on rock crushing blows.
 Generates lots of noise, dust and vibrations.
Classification of mechanical rock drills in accordance to method of
drilling.

Type of Rotary hand Top Hammer Down To Hole P- Turn

drills column heavy duty

Feed force
(Fc)

Rotation
Torque (Mrm)

Percussion
impulse (Fu)

Coefficient of
rock hardness ≤ 10 6 14 8 20 8 20

Drilling Rotary -R R-P P-R P-Turn

method
THE MAIN COMPONENTS OF PERCUSSION ROCK DRILLS

1. The rotation mechanism-

 The percussive-Turn rock drill is equipped with a special rotation
mechanism to enable it causing the rotation of drill bit through the
angle β after every consecutive piston blow forcing the inserts or
buttons to work on a fresh portion of the rock on hole face.
 In order to deliver the maximum power possible, the drill steel is
rotated on the back stroke and not on the fore drilling stroke.
 The main elements of the rotation mechanism are the piston, the
rifle bar with paws and the ratchet.
2. Drilling steel (rod):
 Round – mainly used for drifting and development work
 Hexagon steel – mainly used on the sinker and
jackhammer
 Quarter octagonal section- mainly used on the stopper.
 Percussive- Turn drill steel is made hollow to permit water
and also air to be forced through the steel for washing out
the drill cuttings.
3. Shanks : Is the end of the drilling steel that is struck by the
piston.
Quality of drilling steel and shank:
 Steel should be easy to forge;
 Carbon content should be such that the bit can be properly
hardened through heat treatment;
 The body of drill steel must be tough and stiff enough to
withstand bending and shock;
 The shank must be tough enough to stand up the blows of
piston and avoid excessive wear on its face with the piston.
 Iron, carbon and manganese are the chief constituents of
drilling steel

4. Drill bits:
 The two types of drill bits commonly used on percussion drills
are:
1. The brazed bit;
2. The button bit.
Brazed bits
 The brazed bit is made up from one, four or six rectangular prisms of
cemented tungsten carbide or hardened steel inserts.
 Brazed bits made of 4 prisms of inserts are the most common on
percussive drills and are of two types:
1. The cross bit (+);
2. The x-bit.
 The cross bit is made of inserts which are mounted at 900 from each
other.
 In the X-bit the angle between inserts is 800 and 1000.
 Bits have one hole at the centre and four on the side wall of the bit
for water passage to cool the bit and flush drilled products out of the
hole.

The brazed bit

Button bits:
 Also employ tungsten carbide inserts but their shape is cylindrical.
Advantages of button bits over brazed bits:
 In larger holes there is a more even distribution of cutting elements
over the hole bottom, giving them faster penetration rates.
Disadvantages of button bits over brazed bits:
 They cannot be re-sharpened. However, this disadvantage could
be overcome by using button bits with polycrystalline diamond
compacts (PDC) added on the surface of the button inserts to
increase their resistance to wear and application competitiveness.

Button bit
THE DRILLING INSTRUMENTS OF ROTARY DRILLS

The types of drill bits for rotary drills:.

 These drills have various types of detachable drill bits specific for
drilling on coal and on other soft- medium rocks.
 Roller cone bits (tri-cone ) are the most common bits used for rotary
blasthole drilling
 Bits have three or more cones (rollers or cutters) made with
hardened steel teeth or tungsten carbide inserts of various shape ,
length and spacing .
The types of drilling steels for rotary drills:
i. Steels made of spiral flutes (grooves) for drilling without hole
flashing;
ii. Smooth-hexagonal drilling steels for drilling with hole flashing.
CONTROL OF DRILLING REGIMES- PERCUSSION

 The parameters of Percussion drilling regime

1. The power of piston blow
2. The frequency of piston blow
3. Rate of compressed air supply
 The power of piston blow can be determined by
- Effective diameter of the piston,
- its weight,
- stroke distance
- its driving compressed air pressure during the drilling
stroke (forward piston movement)
 Reducing the diameter of the piston and increasing the stroke
distance or reducing piston weight and increasing the driving
compressed air pressure will increase the power of piston blow.
 Both the frequency and the power of piston blows should be as
great as the drilling steel could withstand.
CONTROL OF DRILLING REGIMES- DTH
Important conditions for the effective DTH drilling process:
1. The correct selection of the drill rig, its instruments and regime are
the main conditions of effective drilling process;
2. Qualification of the driller is also important in the effective
operative regulation of drilling regimes.
The parameters of drilling regime in DTH drilling:
1. Feed Force: - must deliver the effective penetration of the drill bit
into the rock. Feed force increases with the increase of drill
power.
Too high feed force will result in:
- Increased risk of steel getting stuck because of reduced speed
of rotation;
- Increased risk of hole deviation because of flexure (bend) of
the drilling steel.
Too low feed force will result in:
- Reduction of the energy transmitted as the joints tend to
become loose.
2. Rotation frequency- is determined by the energetic
parameters of the piston drill, the geometry of the drill bit
and rock hardness.
- Increasing rotation frequency will increase the mechanical
penetration rate (speed) but bit wear and energy
consumption will also go up.
3. Rate of compressed air supply- Increasing by 2 to 3 times
the rate of compressed air supply could increase
penetration rate by 1.8 to 2 times.
4. The blow power of drill piston –is not regulated parameter.

 The above parameters must be well matched with the

parameters of the piston drill, the physical-mechanical
properties of rock, and the form and size of drill bit.
CONTROL OF DRILLING REGIMES- Rotary- percussive tri-
cone roller bit
 The drilling regime parameters of rotary- percussive tri-cone
roller bit drilling are
1. The feed force on drill bit – P;
2. The rotation frequency of drill bit – n;
3. The quantity of compressed air/ air – water mixture
supplied for hole cleaning and bit cooling – Q
 The quantity of compressed air/air – water (Q) is not a
regulated parameter in rotary percussive tri-cone roller bit
drilling but should be just sufficient for hole cleaning and bit
cooling.
CONTROL OF DRILLING REGIMES- ROTARY

 The regime parameters for rotary drilling are:

1. The feed force exerted on the drill bit;
2. The rotation frequency of the drill bit;
3. Air consumption for hole cleaning (for rotary drilling carried
out with hole flushing).
 Feed force:
- The penetration rate of rotary drilling is mainly determined by
the value of feed force that is exerted on the drill bit
- The feed force per unit area of contact between drill bit and
hole face must be higher than the rock resistance to
compression crushing:
 The rotation frequency of the drilling steel: Increasing rotation
frequency causes the following:
- Penetration rate increases;
- Much higher intensity of power consumption and bit wear;
- Sizes of drilling products become smaller
 Air consumption for hole cleaning : determined by the
following:
- The granulometric composition of drill products;
- The intensity of dill product generation/penetration rate.
THE TYPES OF ROCK DRILLS FOR UNDERGROUND MINE
DRILLING

 These drills are either hand held or machine/crawler mounted.

 Handhelds: the jackleg and the stopper.
 Machine mounted - the drifters/jumbo.
 A jackleg is a jackhammer with hydraulic/pneumatic cylinder
mounted to it to support the drill and to provide the required feed
force during the drilling process. They are used in small drifts,
headings and in stopes (for production drilling)
 Stoppers: They are jackhammers with hydraulic/pneumatic cylinder,
rigidly attached to it. They are used for drilling up-holes.
 Source of power used is either water(hydraulic drills) or compressed
air (pneumatic drills) or electricity.
 Most of underground rock drills employ the percussive power,
rotation torque and feed force.
 Underground rotary drills used for drilling blast holes in coal mining
and in soft rocks of hardness (f ≤ 4)
 The power source for most rotary drills is electricity, which is 6 10
times more cheaper than compressed air.
 Percussion Drilling Rigs

• Rigs are either truck or track-mounted. Rubber tired rigs can travel
quickly between job sites. However not able to move on rough terrain
• Track or crawler mounts can easily traverse rough terrain.
Rotary Drilling Rigs
Jackleg
 Topic 2: FIELD CONTROLS FOR SAFE,
EFFICIENT BLASTING
MAJOR FACTORS INFLUENCING BLAST EFFICIENCY
1. Geological Effects
2. Optimum Explosive Performance
3. Quality control
4. Communication

A. Geological Effects
 Blasting results are influenced more by rock properties and structure
than by explosive properties
1. Physical rock properties- e.g. compressive strength, tensile
strength, Young's modulus, density etc.
2. Rock structure-
 Rock fragmentation is primarily controlled by the rock structure
(i.e. bedding, jointing and faulting)

Adverse bedding planes rock

Massive rock

 Soft or weak bedding planes can reduce energy confinement

and often require the use of decking to achieve optimum blast
performance
 Highly jointed or fractured rock requires relatively less explosive
energy to obtain good fragmentation
 Deck loading
3. Rock hardness
 Can be made to define the relative strength of the rock
 Harder, higher density rocks usually require higher energy
factors for optimum fragmentation unless they are highly
bedded and jointed
 Blastholes should be loaded according to zones of hardness.
4. Reactive ground
 Nitrate based explosives react with sulfides in rockmass to
generate heat (often in excess of 650°C) and toxic gases(
SO2, CO, H2S)
 A common reactive ground type is black pyritic shale
 Excessive heat levels can cause premature detonations or
failure of initiation systems
 Don’t load hot holes over 500C in reactive ground
 Load sulfide rich and or warm holes last, preferably with an
inhibited explosive that will slow the reaction between the
nitrates and the sulfides
 If inhibited explosives can not be used use
temperature resistant liners to separate the
explosive from the sulfides in reactive ground.
 Fire blastholes the same day as loaded
 Beware of misfired holes - they can self
detonate if the adjacent ground reacts after the
blast

5. Water
 Presence of water has major influence on the
type of explosive used and overall costs
 Mine site dewatering - expensive but can be
justified by overall savings in mine equipment
and explosives used.
B. Optimum Explosive Performance

1. Explosive energy distribution in the rockmass

- Energy must be evenly distributed to achieve uniform
fragmentation
2. Explosive energy confinement
- Explosive energy must be confined enough after detonation to
establish fractures and to displace material
3. Explosive energy level
- The energy level must be sufficient to overcome the structural
strength and mass of the rock and while providing controlled
displacement.
C. Quality Control
 Paying attention to details is the key to quality
control and achieving safe, consistent, efficient
blasting.

 Procedures for the design and implementation of

blast designs should be defined, documented,
followed, and audited to insure quality control.

D. Communication
 Safe, optimized blasting requires good
communication between members of each group
and interaction between groups
 Efficient blast designs require a group
effort
I. SITE EVALUATION
Site Conditions That Influence Blast Design are
1. Rock type
a. ore / waste
b. block size
c. structural orientation
d. blastability (easy, average, hard)
e. reactive ground
2. Desired fragmentation
a. ideal size distribution for excavation
3. Desired amount and direction of rockmass displacement
a. dilution control
b. excavator efficiency
4. Wall damage control
a. blast location with respect slope
b. slope stability
5. Water conditions
a. water table
b. recharge rate
6. Available explosives
a. density
b. energy
c. water resistance
7. Drilling equipment
a. type available (rotary, hammer)
b. bit diameter
c. angle capacity
d. productivity
8. Labor requirements
a. productivity (holes loaded per day)
9. Bench restrictions
a. free face
b. berm location
c. overbreak from adjacent blasts
II. BLAST DESIGN
Considerations are
1. Performance goal(s)
2. Site conditions

III. BENCH PREPARATION

1. Bench should be designed to accommodate drill pattern
2. Bench should be level to aid in drill set up
3. Surface runoff water should be diverted away from bench
4. Avoid uneven free faces
5. Train operators and supervisors on the need for clean
straight free faces
IV. PATTERN LAYOUT
1. All blasthole locations should be precisely located
2. Mark location of blastholes with flags or painted rocks (unless
drill guidance is available)

GPS assisted blasthole layout

V. BASIC DRILLING
A. Drill setup and drilling
1. Setup precisely on blasthole marker
2. Position drill as level as possible
3. Avoid over drilling
4. Drill pattern in a logical sequence so pattern can be
squared off and shot when needed
B. Drill record
1. Label each hole depending on its location in the shot
2. Record levels of different hardness by depth
3. Record the length of broken material at collar
4. Record any voids encountered by depth and length.
5. Date drilled
6. Driller's name
7. Measure and record total depth of hole after drill has
pulled away from collar .
8. If the bench is broken case top of hole to prevent unwanted
backfill.

PVC collar reinforcement

9. Record total time required to drill entire pattern

10. KPI –Key performance indicator (KPI) designed vs
actual hole depth , designed vs. actual burden spacing, face
burden, hole deviation collar condition, redrills required, drill
productivity (m /shift), drilling cost ($/m).
Drilling blastholes
VI. BASIC LOADING PROCEDURES
A. Pre-load check
1. Check blasthole depth with measuring tape or pole
2. Check and record water level
3. Compare current depth to depth desired
4. Redrill short holes if possible
5. Backfill overdrilled holes
6. Calculate the amount of explosives needed for both wet
and dry holes
7. Check hole temperature in reactive ground.
8. KPI’s – actual vs desired hole depth, number of wet holes,
water depth, lost holes.
Backfilling overdrilled holes
B. Explosive quality control
1. Check density of bulk products at least once per truck
load
2. Check expansion of gassed products prior to stemming

Checking bulk product density

C. Blasthole loading
1. Assemble primers immediately before placing in blasthole
2. Load holes according to hardness zones and water
conditions
3. Check explosive column rise with measuring tape to avoid
Overcharging
4. Use bagged explosives or liners in zones where
overcharging can occur
5. Avoid contamination of bulk explosives with drill cuttings
while loading
6. Record specific hole conditions, loading and stemming for
each blasthole
D. Blasthole stemming
1. Use angular sized material for stemming approximately
1/10th the blasthole diameter
2. Longer stemming lengths should be used when stemming
with drill cuttings
3. KPI – stem length range vs design.
Loading booster and detonator into Blastholes
Loading explosive into Blastholes
E. Connecting blastholes for initiation
1. Remove all equipment from blast area.
2. Allow only qualified personnel on shot to perform tie-in.
3. Assign responsibility for tie-in to one individual
4. Record the delay layout and timing configuration
5. KPI – completed timing diagram
VII. BLAST GUARDING AND BLAST DETONATION
1. Blaster-in-charge should follow blast clearing procedures
using an established form or check list
2. Blaster-in-charge responsible for removing all equipment and
personnel from blast area.
3. Blaster-in-charge assigns guards to block access to blast
zone – note guards should understand their authority and
responsibility.
4. Fire blast as soon as possible after tie-in has been completed
5. Set up remote videotape for blast recording
6. Record the time of detonation
7. KPI – completed pre-blast clearing procedure check list
Understanding
Warning Signals &
Signs
POST BLAST CLEARING
1. Wait until shot gases clear before entering the area.
2. Misfire detection
- listen during detonation
- examine muckpile profile-damaged signal tube, electric
detonator wires, detonating cord, or safety fuse.
- oversize material
- explosives in muckpile
4. Potential misfire causes
- initiation system damage or improper hook up
- inadequate primer or improper primer location
- malfunction of explosives
5. KPI – number of misfires
VIII. POST BLAST REPORTING AND ANALYSIS

A. Shot reports-require regulatory information (e.g. date of blast,

time of blast, pattern description etc.)
B. Post shot analysis
1. Determine the performance of the blast qualitatively in terms
of goals( fragmentation, flyrocks, wall damage, muckpile
displacement).
2. Review videotape to qualify blast performance in terms of:
► face and surface movement
► stemming confinement
► source and magnitude of flyrock
► source of oversize material
► backbreak
► potential misfires
3. Calculate drilling and blasting cost
IX. EXCAVATOR PERFORMANCE ANALYSIS
1. Quantify performance with excavator cycle time evaluation
2. Determine if oversized material came from the interior or the
exterior of the shot
3. Report unshot explosives uncovered during excavation to
supervisor
X. BLAST DESIGN REFINEMENT
A. Fragmentation improvement
Common causes of poor fragmentation or difficult excavation
are
1. Improper energy distribution caused by:
- too large charge diameter
- disproportionate burden and spacing
- too much stemming
- blasthole deviation
- excessive face burden
- adverse geology (such as open jointing)
2. Improper energy confinement caused by:
- poor stemming material (drill cuttings)
- incorrect stem length
- adverse geology (weak seams, voids)
- wrong blasthole pattern for charge diameter
- unsuitable delay configuration
- excessive energy level
3. Improper energy level caused by:
- incorrect pattern or charge diameter
- unsuitable delay configuration
- insufficient water resistance of explosive
- bad explosive quality control
- inappropriate primer location
B. Controlling air blast and vibrations
1. Ground vibration control
- provide relief
- accurately layout and drill blasthole pattern
- reduce the weight of explosive per delay by reducing hole
diameter, bench height, or decking the powder column but
maintain relief.
- increase delay periods in back rows
- direct propagation away from nearby structure
- use accurate delays
2. Airblast control
- confine charges (flyrock and airblast go hand in hand)
- proper stemming length
- accurate drilling
- proper face confinement
 - don’t load weak seams
- if possible avoid using detonating cord
- avoid blasting in adverse weather conditions (wind
blowing toward point of concern, etc.)
- slow delay interval down face to more than 1 ms /ft of
burden
3. Controlling Flyrock
4. Wall Control
MISFIRES
 The most hazardous part of blasting is the handling of misfired
explosives.
 When a misfire does occur, the degree of hazard depends on
the type of explosive used in the blasthole.
 When a misfire occurs, the power source used to initiate the
blast must be disconnected, the firing line shunted or made safe
before entering the blast area to inspect the misfire.
 All personnel must stay out of the blast area for at least 1/2
hour. Access to the blast area must remain blocked and
guarded.
 Proper misfire handling should be conducted by experienced
individuals familiar with the initiation systems and explosives
used, as well as the proper techniques to handle, neutralize and
render safe the explosive materials.
 In the general case, if the blastholes are relatively intact, the
explosive should be shot in place (refiring)
 In cases where the rock is cracked near the blasthole, it may be
necessary to pile dirt or soil on top and around the hole to
ensure that flyrock may be reduced or eliminated.
 If the initiator has fired or has been tom away from the
hole, the next option would be to remove the stemming
and reprime the hole.
 The stemming could be removed pneumatically with
compressed air or possibly with water pressure and a
new primer placed on top of the main explosive column.
 If a primer is placed on top of the column charge, it can
be stemmed and fired.
 Do not permit any work in the misfire area
 With proper careful blasting procedures, the instances
of misfire are rare
 Refiring a misfire is usually the safest and best way to
eliminate the danger.
STORAGE OF EXPLOSIVES

 Explosives should be stored in accordance with federal or

state and local regulations
 When explosives are stored in a magazine, the magazine
should be clean, dry, cool, well-ventilated, and properly
located.
 The magazine should be constructed so that it is bullet-proof,
fire-resistant, and meeting all federal or state codes.
 Initiators such as electric blasting caps, Nonel, or any other
type of blasting cap should not be stored in the same
magazine with high explosives.
 Explosives in magazines should be thoroughly marked as to
the date of purchase and the oldest products should be used
first
 No source of fire flame should be brought near an explosive
magazine, and the magazine should be located so that there
is no grass, brush, or debris nearby.
 When explosives are brought to the job, they must be stored
in day boxes that meet federal or state codes.
 They should be placed in an area that is not in any danger
from falling objects, fire or heavy equipment

Storage of Explosives
TRANSPORTATION OF EXPLOSIVES
 When explosives are transported on the highway, they
should be transported in vehicles in proper working
condition and equipped with federal, state or locally
approved containers for safe transport
 Unless the explosives are in the proper approved
containers, caps and explosives should not be carried
on the same vehicle.
 The explosive cargo should be gently unloaded and
cases should not be thrown onto the ground.
Transport of Explosives