Unit 1 (LO1)

o Definition: Surface Mining is also known as “Strip Mining”. It is a system of extracting coal or other valuable
minerals, by removing the layer of soil and waste, deposited over it, called the “Overburden”, which exposes
the layer of mineral for Extraction. The extraction in this method, is carried on by making benches in the
area. This method is useful for mineral deposits that are located at a shallow to moderate depth form the
surface.

The basic advantage of this method is that there is

no requirement of roof support etc. & there is no provision of ventilation, as it is done by the open
atmosphere surrounding it. Moreover, the percentage of extraction of mineral is also comparatively high at
around 80-95%. Due to all of these advantages, the Surface mining method is generally preferred over its
counterpart, i.e., Underground Mining. However, this method is feasible, only when the thickness of the
overburden material is less.

o Terminology: some major ones are as follows:

▪ Bench: A bench is a ledge, which forms a single level of operation, where ore & waste are excavated.
In other words, “An exposed rock/mineral block separated by upper and lower surfaces is called the
Bench.”
▪ Bench Height: The vertical distance between the upper and lower surfaces of each bench is called
the Bench Height.
▪ Bench Face: The exposed sub-0vertical surface of a bench is the bench face.
▪ Face Angle: The average angle that the bench face makes with the horizontal is called the Face
Angle.
▪ Bench Floor: The exposed bench lower surface is the bench floor.
▪ Bench Width: The distance between the crest and the tow measured along the upper surface is
called Bench Width.
▪ Bank Width: The horizontal projection of the bench face is the Bank Width.
▪ Cut: The width being extracted from the working bench is called the Cut.
▪ Width of the Working Bench: The distance from the crest of thee bench floor to the new toe
position after the cut has been extracted is called the width of the Working Bench.
▪ Safety/Catch bench: After the removal of cut the remainder of the bench is called the safety bench
or cut bench. Purpose of safety/catch bench is to collect the sliding rock boulders from top benches
and to arrest them. This type of benches are left on every level. The width of such benches varies
from 65% to 70% of the height of the bench. However, at the end of the life of the pit, it reduces tot
30% to 35% of the height of the bench.
 ▪ Berm: Piles of broken rock, material constructed along the crest to improve mine safety is called the
Berm. It acts as a guard rail to prevent trucks and other machines from backing over, arrest
rolling/broken rock boulders, control of noise, etc. Height of the berm is generally kept higher then
the radius of the tyre of the truck/dumper.
▪ Trench: A trench is a type of excavation made in the ground, that is generally deeper than its width,
and narrower compared to its length.
▪ Ingoing Trench: Leads inside the mine
▪ Production/Working Trench: the place where from extracted minerals are find out
▪ Overburden: also “waste Spoil”. It is the material that lies above an area that lends itself to
economical exploitation, includes rock & soils that lies above the ore body.
▪ Inter-burden: It means the waste material that lies in layers between ore deposit. It can also be
called “Gangue”. It includes the parting that separates the seams in a mining field.
▪ Cut-Off Grade: The cut-off grade is the level below which, material within an ore-body does not
contain sufficient value to economically justify processing it into a final form.
▪ Box-Cut: Initial cut at the start of quarrying or bench making.

o Classification: Can be done on various bases, major ones among which are as follows:
▪ On the basis of Method: is as follows:
▪ Open Cast Surface Mining: When the mineral body is not thick, but is extended over a large
geographical area, the Opencast Surface Mining Method is suitable. The major speciality of
this method is the process of “Casting”. In this process, the area that has been emptied,
after the removal of O.B. & Mineral, is filled/casted with the O.B., that is removed from the
successive stage or strip of land. In other words, the O.B. is not transported to waste dumps
for disposal, but is casted directly into the adjacent mined out area.

This process continues for the entire mining operation, and in this
way, the Reclamation continues along with the Exploitation/Extraction Stage. Material
Handling thus consists of excavation of casting, generally combined in one unit operation
and performed by a single machine.

In the Opencast method, normally different equipment is employed for the purpose of
Stripping O.B. & mining coal.

# Advantages: are as follows:

Productivity
1. Highest productivity in all surface mining methods Production rate
2. Lowest cost per tonne of coal produced (relative cost 10%) Production time
Cost per ton coal
3. High production rate & Early Production Labour
4. Low-Labour Intensity & Blasting Cost Blasting cost
5. Relatively Flexible & Suitable for large equipment Development
6. Simple Development & Access. Flexible
7. Normally eliminate the requirement of haulage of O.B.

# Disadvantages: are as follows:

1. Limited Depth (90 m.)

2. Limited Stripping ratio (1.3 to 1.9 m3/tonne)
3. Surface is damaged, extensive environmental reclamation is required
4. Skilled operators are required
5. Public image is negative
6. Weather can impede operations
7. Slope monitoring is required to be monitored & maintained.
▪ Open Pit Surface Mining: Open Pit mining is the process of mining any near surface deposit
by means of a surface pit excavated using one or more horizontal benches. The major
difference between his and the Opencast method is that the “Overburden is deposited at a
separate place during the entire mining operation.” In this way, the process of reclamation
begins after the completion of the process of Extraction of mineral.
This method is suitable when the mineral body is not spread over a
large area, but is confined to a small area. Another difference between the two is that the
horizontal span of this method, is relatively small.

# Advantages: are as follows:

1. High Productivity & production Rate (OMS 100-400 tons)

2. Relatively Lowest cost
3. Low Labour Requirement, relatively unskilled labour can also be employed
4. Ideal for large equipment, also flexible in operations
5. Favourable health & safety factors, No U/G Hazards
6. Relatively less area of land wasted

# Disadvantages: are as follows:

1. Limited by depth (around 300 m.)

2. Limited by Stripping ratio (0.8 to 4 m3/tonne)
3. High capital investment associated with large equipment
4. Reclamation is additional, hence huge costs involved
5. Slope Stability is critical to all operations
6. Pit may get filled with water after mining, which poses threat.
▪ Quarrying: It is a special method of Surface Mining, that does not use drilling and blasting
for extraction of mineral, but employs large scale cutting equipment, for cutting out large
blocks of minerals. This method is not employed for mineral that are relatively hard and
valuable, such as Marble etc. Its suitability also requires the thickness of Overburden to be
relatively small.
The difference between it and the surface Pit method, is that the benches (called
faces) are lower and generally vertical. This method is highly selective, small-scale method
with low productivity. It is also the costliest method, and is used only when other methods
are not feasible.
# Advantages: are as follows:
1. Low capital cost, non-extensive mechanisation
2. Suitable for small deposits and small-scale mining operations
3. Easily accessible
4. Stable walls and benches, generally n bank support required.
5. High selectivity, can discard low quality stone.
6. Good safety, low chance of slope failure.
# Disadvantages: are as follows:
1. Somewhat limited by depth (90 – 300 m.)
2. Low Productivity, High Labour cost as requires skilled labour
3. Highest Mining Cost because of low productivity (relative cost 100%)
4. Inflexible, mining plan cannot be easily changed at depth
5. Mechanisation is limited
6. Complicated & costly rock breakage, because of the inability to use full power of
explosive
7. High waste percentage (60-90%)
▪ Mountain Top Removal: It is also called “Contour Mining”. This method is commonly
employed for deposits located in hilly & mountainous terrain.
 In India, most of the Iron-ore & Bauxite deposits are located in hilly terrains
& they are generally worked with Contour Mining Method.
▪ Highwall Mining: It is a remotely controlled mining method, which extracts coal from the
face of a coal seam, under a highwall in a surface mine, which has reached the final highwall
position, due to Uneconomic Stripping Ratio. The coal left in the high-wall can be extracted
by this technology, which otherwise would be lost for ever. SECL was the 1st company to
implement this technology in India.
▪ On the basis of Mechanization:
▪ Manual Mining: It is a method of mining in which, most of the tasks and operations are
done by labours and manpower. In these mines, small drill equipment is used. The drill holes
are 1.2 to 1.8 meter deep and 37 mm. in diameter. Gun powder or other low explosive are
used for blasting. The broken cool and overburden, is loaded onto the tubs, by the labourers
and is then transported out by the haulage engine.
▪ Semi-mechanized Mining: In this method, the operations are divided amongst the labours
and small machines. There are only a few mines of this type in the country. In these mines
also, the blasting is done by Gun Powder and other low explosives, but the broken material
after blasting, including the overburden and coal, is picked up and loaded by small machines
like bulldozers etc. Similarly for transportation, trucks and dumpers are used.
▪ Mechanized Mining: In this system, heavy machineries, like drag lines etc. are used for
removal of overburden and coal. Also, well-bore drilling machine are used for drilling holes
for explosion. The drill holes are 5 to 18 meter deep and 125 to 300 mm. in diameter. High
explosives like Liquid Oxygen & opencast gelignite etc. are used for explosion. The
transportation of broken material is done by loco engines, belt conveyor or by big trucks
called dumpers.
▪ On the basis of Machines used:
▪ Dragline: A Dragline is a costly machine, that is used for the removal of overburden only. It is
used in the following conditions:
▪ Where the thickness of overburden is 50 m. or less
The larger is the bucket and the longer is the boom of the
machine the greater is the efficiency of the machine.
▪ Where the seam is thinner than 25 m. or where the parting between the two seams
is less than 25 m.
▪ The seam that is completely free from any fault or other geological disturbances
▪ For a large extraction area, where there is only a single seam or two seams with
small parting, as the machine cannot travel greater distances.
▪ Shovel Dumper: It is the most popular combination and can be used in any geological
condition. In this combination both the overburden and the coal are extracted. The loading
is done by the shovel and the transportation is done by the dumper.
▪ Pay Load Dumper: It is also called “Wheel Loader” and it is a loading machine that is
mounted on wheels. It is suitable for loading of broken coal and is not useful for loading of
overburden.
▪ Shovel-Conveyor: In this method, the broken coal reaches to the Mobile crusher from the
shovel and it is then transported to the surface, by the help of conveyors. The overburden is
separated and sent to the overburden yard. In India, it is used in only used in only one
Opencast Mine.
▪ Scraper: This machine is only used to remove the soil at the surface, which is unloaded at a
separate place at some distance. It is not used for the breaking and removal of coal.
▪ Dozer Ripper: It is basically a dozer, that has a plough shaped attachment at its rear portion,
which is called a “Ripper”.
A ripper is a hydraulic operated device that rips the ground, in form
of long strips. It is a useful equipment in case of a seam, which has a band of stone or a
parting between the thick seam. In such cases, the use of shovel-dumper is not economic
and also disrupts the production
 ▪ Bucket Wheel Excavator: This machine is only useful for the cutting and removal of soft soil,
lignite or other similar soft stones. In India, it is only used for lignite mines which is covered
by a thick layer of soft soil.
▪ Surface Miners: It is also a coal cutting machine, that is used to cut thin layers (0.2 to 2.0 m)
of coal and load them upon Dumpers and Belt Conveyors. It is also useful in cases where a
thin band of stone is present in between the seams
▪ Hydraulic Monitor: It is a remote-controlled device, that throws a high-pressure jet of water
and it is used for the removal of the layer of overburden, deposited over a coal seam. In the
case of a mountainous terrain, the overburden gets flown along with the stream of water
and the coal seam gets exposed.
▪ Combination: In India, the open cast mines are worked in either of the two manners:
In the first type, the operations of the removal of overburden and the removal of broken
coal are done by the Shovel-Dumper only.
On the other case, the work is done in combination as:
1. Upper overburden by Shovel-Dumper
2. Lower rocks by Dragline nature
3. Broken coal by Shovel-Dumper Min.Depth
Physio mechanical
The manner that has to be used, depends on various factors like; Surface condition
Surface structures
1. Extent of excavation area Climate
2. Number of seams oms
Stripping ratio
3. Presence of Dirt bands Req production
4. Inclination of seam. Mechanization
Capital
Coal quality

o Factors affecting Choice of Opencast Method: are as follows:

▪ Nature of Deposits: Under this, the geological method and the shape of the formation of the deposit
is considered.
Generally, if the deposit is at a very less depth from the surface, than it is feasible to
perform extraction with Opencast Method. Similarly, if the seam is very thin and the cost of removal
of Overburden is great, it is not feasible to work with Opencast system, as the initial cost would be
too high.
▪ Minimum Depth of Coal from Surface: Inn case of a deposit, when the minimum depth is large and
the stripping ratio is also small, the feasibility of Opencast system ends. It is also somehow
depending upon the quality of coal.
▪ Physio-Mechanical Properties: Weak strata & the strata with geological disturbance are difficult to
be worked with underground method & hence, Opencast method is preferred in such cases.
▪ Stripping Ratio: A stripping ratio is useful in determining, whether the mine is economic. It is the
ratio of the quantity of coal produced, to the amount of overburden removed. If the Stripping ratio is
greater, the mine is not considered economic.
▪ Required Production: The amount of Output required is useful in determining the layout of benches,
No. of overburden benches, and the age of mine etc. In order to achieve high output, it is necessary
that high efficiency machines are used for the breaking and transporting of Overburden and Coal.
▪ Degree of Mechanization: The Layout of Bench, Number of Overburden Benches etc. also depends
upon the degree of mechanization. In addition, the machineries used for drilling, loading &
transporting shall be of high efficiency.
▪ Available Capital: During the removal of Overburden, Huge Capital Costs in terms of HEMM and
manpower are required, but the production is zero. The available capital shall be large for an
opencast mine.
▪ Surface Conditions & Climate: The choice of a proper ground surface affects the selection of Proper
dumping Ground, Haul Road of proper Gradient, Discharge of water from pit bottom, Stock yard, etc.
 A place with excessive rains, are not suitable for Opencast Method. Similarly, A place that has
excessively hot or coal weather, is also unsuitable for Opencast Method.
▪ O.M.S.: A place, where other conditions are suitable and a high O.M.S. is required, the Opencast
Method is employed.
▪ Quality of Coal: The greater is the quality of coal, the greater will its price be. Hence, even if the
production and operation costs are higher, the opencast method is suitable until the quality of coal
is sufficiently high.
▪ Surface Structure: For the opencast method, the place with no surface structures and residence are
selected, because of the high cost of compensations, in the case of Residential areas.

o Advantages: are as follows:

▪ Ventilation: In the opencast method, there is no need of Ventilation. The harmful gases from the
seam are immediately diluted in the general air.
▪ Mechanization: Huge scale machines are used in Opencast method, due to which, the rate of
production and the speed of production are sufficiently fast.
Ventilation ▪ Drilling & Blasting: In opencast method, the blastholes are 20-200 mm. din dia. And are filled with
Lighting around 20-25 kg explosives, at a time. This contributes to a comparatively large production rate.
Roof support ▪ Roof Problem: Surface Mine has no roof. This entirely removes the need of Roof Support. There is
Training
Supervision no need of Highly detailed monitoring of Roof as well as the huge costs involved in supporting it.
Drilling blasting ▪ Lighting: In an opencast mine, the task of Lighting is done naturally during the day time, by the sun
Transport and by means of Large Mercury Lamps in the Night time.
Mechanization ▪ Transportation: As there are sufficient roads in the surface mines, the material transportation is
Large production
done by big trucks, called Dumpers. Also, due to huge large space available, Conveyors can also be
used for the purpose.
▪ Training & employment: In a surface mine, the training of workers is comparatively easier. Also,
female Workers can be employed here.
▪ Supervision: The task of supervision is also comparatively easier in the surface mines.
▪ Extraction: The percentage extraction in the surface mines is as high as 80-90%. The O.M.S. is also
comparatively higher.
o Disadvantages: are as follows:
surface destroyed ▪ Weather: The working in a surface mine, during the night in winter season and the afternoons of
Large area for disposal
large capital
summer season, becomes slower due to extreme temperatures. Also, During the rainy season, if
Huge machinery there is no provision of removal of water, the production comes to a halt, due to storage of water in
More pollution the pit bottom.
Surface vibrations ▪ The surface gets destroyed completely. The agricultural land, if any, gets distorted completely.
weather
not for deep seams
▪ If the depth of the coal seam is greater, there is no profit in extracting it with surface methods.
▪ In comparison to Underground methods, the Surface mining has higher level of pollution.
▪ A large area of land is required for the deposition of mine tailings and other wastes, extracted during
the Mining Operations.
▪ A large capital is required for the removal of Overburden prior to any production.
▪ During the removal of Overburden, the O.M.S. becomes almost zero.
▪ The overburden may get mixed with the coal, which reduces the quality of coal.
▪ If the Overburden is carbonaceous, or if it contains some quantity of coal, there is a risk of fir in the
overburden dump.
▪ During the blasting, drilling and other operations, there is a large quantity of Dust in the surrounding
atmosphere. If water is not used for dust suppression, then it affects the atmosphere seriously.
▪ Very High-Power Explosives are used for blasting operations in surface mines. This can cause damage
to nearby structures and residential buildings. Hence, a surface mine is established at a distance of
at least 300m. form any residential area.
 o Stripping Ratio: In order to extract the coal from a seam, using the surface methods, the overburden is
required to be removed. The ratio between the quantity of coal produced to the quantity of overburden
removed, is termed as “Stripping ratio”.
𝑞𝑢𝑎𝑛𝑡𝑖𝑡𝑦 𝑜𝑓 𝐶𝑜𝑎𝑙 𝑃𝑟𝑜𝑑𝑢𝑐𝑒𝑑 (𝑖𝑛 𝑡𝑜𝑛𝑛𝑒𝑠)
𝑠𝑡𝑟𝑖𝑝𝑝𝑖𝑛𝑔 𝑟𝑎𝑡𝑖𝑜 =
𝑞𝑢𝑎𝑛𝑡𝑖𝑡𝑦 𝑜𝑓 𝑜𝑣𝑒𝑟𝑏𝑢𝑟𝑑𝑒𝑛 𝑟𝑒𝑚𝑜𝑣𝑒𝑑 (𝑖𝑛 𝑚3 )
The greater is the stripping ratio, the greater is the cost of removal of Overburden, and the lesser will be the
productivity of the mine.

▪ Characteristics: is as follows:
▪ Gradient of Seam: Due to the inclination of seam with the horizontal, the depth of the seam
will continuously increase in the dip direction. As a result, Stripping Ratio will also increase.
The greater is the inclination, the greater is the increase of the Stripping Ratio.
▪ Change in Thickness of Seam: The stripping ratio will be less, where the seam is thick and
vice-versa.
▪ Topography of the region: In the case of mountainous terrain, the Stripping ratio changes
frequently. In the direction of peak, it will be great and less in the direction of valley.
▪ Multiple coal Seams in the region: The stripping ratio can also change due to this reason.
The stripping ratio keeps increasing in the direction of dip of a seam. It suddenly decreases,
on the founding of another seam and continues doing so.
▪ Parting Thickness: In case of a place, where the thickness of the seam changes frequently,
the Stripping Ratio also changes accordingly.
▪ Effect of Geological Disturbances:
i. If the seam has shifted down due to fault, the stripping ratio will increase. Similarly, if it
has shifted up, the stripping ratio will decrease. A fault can also make an area, a coal
Free Zone.
ii. Wash Out Region can either increase the Stripping Ratio or make the area, a Coal Free
Zone completely.
iii. The seam Splitting can also heavily change the Stripping Ratio.
▪ Types: are as follows:
▪ Maximum Allowable Stripping ratio: is as follows for different categories of mines:
i. Mechanized Mines: The Stripping ratio for these shall not be greater than 1:4
ii. Semi-Mechanized Mines: The Stripping ratio for these ranges between 1:6 to 1:8
iii. Manual Mines: The Stripping ratio for these shall not be greater than 1:10
▪ Overall Stripping Ratio: The ratio between the entire amount of coal present in the area, to
the Total amount of overburden required to be removed, is known as Overall Stripping ratio.
▪ Break Even Stripping Ratio: The Stripping Ratio of a point of a surface mine, where there is no
overall profit in extraction, which means that the cost of the removal of overburden, will be equal to
the value of coal obtained, is known as Break-Even Stripping Ratio. This condition is referred to as
“break-even”. The Mining operations are not continued beyond such point, as it would lead to
losses. The break even stripping ratio,

To be noted more from bookmarks and notes

o Elements of Benches: The major elements of the benches made in an Opencast Mine, are as follows:
▪ Height: The height of the bench in an opencast mine, depends upon the machine used for the
removal of the Overburden, in that mine. In case of a manual mine, the height of the bench, is kept
between 1.5-3m., and that in the Mechanized mine, it depends upon the Boom Length of the
machine used.
The height of the benches in the mine, is also depended upon the
nature of the Overburden. If it is soft, the height should be kept less, in order to prevent any
accidents, due to the slide of benches. But if the O.B. is hard, the height of the benches can be kept
large.
 Subject to prior special permissions, the height of the Benches, can also be kept up
to 1.2 times the length of the longest boom of all machineries used in the mine.
If the height of the benches is increased, then there will be less no. of benches, and it will
benefit the efficient Loading & Transporting operations. There will be less requirement of Levelling
work on the top of the benches.

Longest Boom Length * 1.2 = Height of Bench

# Factors affecting Width & Height of the Benches: are as follows:

a. Provisions: According to the provisions of the rules and regulations, the height of the benches
shall not exceed the length of the boom of the machine used in the mine.
b. Thickness of the Deposit: The height of the bench, also depends upon the thickness of the seam.
c. Nature of Strata: Under this, the sliding friction, and the sliding angle, made between the various
layers of the overlying strata are considered. If the overlying strata is hard, the height of the
bench can be kept more. But if the strata are broken, the height has to be kept low, to prevent
accidents.
d. Water Pressure: the water present in the strata exerts a pressure on the sides of the benches.
Due to this, there is high possibility of bench failure, caused due to its sliding by the water
pressure. Hence, in places where the water pressure is high, the height of the bench is kept low.
e. Geological Disturbances: The fault Planes & the weak Bedding planes, passing along the corners
of the benches, poses a great danger on the stability of the Benches, hence, in such conditions,
the height of the benches, is kept low.
f. Equipment: It is an important factor, in determining the height of the benches, in an opencast
mine. In places, where Deep Hole Blasting is done, the height of Bench is confined, to 80% of the
capacity oof the machine that is used to drill the holes. Shovel is used to remove both coal and
O.B., but the dragline is used for the removal of O.B. only, Hence, in case of places, where
Dragline is used, the height of the benches, is kept equal to the length of the Boom.

The height of the bench also affects the transportation system. If the
height of the benches is more, then the gradient will also be more and there will be more problems,
in the transportation of material.

# Advantages of High Benches: are as follows:

g. Efficiency of shovel is increased, if the bench height matches the height matches the machine
height.
h. It minimises the amount of ripping and levelling out of the berm.
i. Efficiency of the transport system is improved because of lesser no. of benches.

# Disadvantages of High Benches: are as follows:

j. In case of presence of joints and slabbing, the probability of Slope Failure is increased.
k. Because of Boulder falling from top, severe damages of machinery and workers can happen.
 l. Supervision of high benches is difficult.
m. Deep Hole blasting in such places, may yield oversized fragmented rock, which may require
secondary blasting & also require heavy amounts of explosive.
n. Complexity may arise during selective mining.

▪ Width: The provisions regarding the width of the benches, are as follows:
o. The width of the benches, shall not be lesser than the height
p. If the dumpers have to be operated, the width of the benches, shall not be lesser than 3 times
the width of the benches.
q. The width shall not be lesser than the width of the widest machine to be used in the mine + 2m.
r. The width shall be the widest dimensions, obtained from all of the above provisions.

Where blasting is done of a row, the width is kept low, and blasting is done
in Multiple rows, the width is kept large.

The width also depends on the manner oof blasting & the transportations system. In places, where
rail Transport is used in place of Truck Haulage, the Bench Width is kept greater, than that in the
Truck Haulage system that has larger manoeuvrability & flexibility of operations.

# Factors affecting the Width: are as follows:

s. The width of the muck pile

t. Width of the dumper movement
u. Width of double/single lane traffic
v. Clearance for safety of truck from edge of benches.

The bench width, if not optimum or excess, directly adversely effects the
economy of the mine. After reaching to the limit of pit, the width of the bench is generally reduced
for giving a shape of overall required pit slope.

▪ Angle of Slope: The Angle of Slope mainly depends upon the Angle of Repose. The angle of Repose is
the steepest angle at which a sloping surface formed of Loose material is stable. In other words, the
angle of maximum slope at which, a heap of any loose solid material (as earth) will stand without
sliding, is known as the Angle of Repose. The Angle of Slope on the mines, is kept less than the Angle
of Repose, in order to prevent the accidents in the mines.
Each type of material or strata has its own Angle of repose. This angle differs, when in Loose
material or strata. At the Angle of Slope, the Slope Bench does not slide on its own and the bench
remains stable. The risk of bench failure, or sliding is less, if the Angle of Slope is kept proper. The
Angle of Slope depends upon the following factors:
w. Slope Geometry
x. Geographical Conditions
y. Surface Geology
z. Water quantity in the strata.
aa. Method of Mining
bb. Time Available
cc. If the working is old or new.
dd. Plane of weakness
ee. Orientation of the bedding plane.

# Slope Failure happens mainly due to:

ff. Slow process of rational shear.

gg. Mechanical properties of rock, like:
i. Angle of Friction
ii. Cohesion
 iii. Pore Water pressure
iv. Seepage Forces
v. Tension Cracks etc.

In order to protect the bench from failing, the pressure has to be reduced, and the bench has to be
made stable. In the case of the sedimentary stones, the Angle of Slope is kept between 50-60
degree. It is kept less in case of a strata with water. In competent rocks, angle of slope of the bench
varies between 70-85 degrees. This slope may be obtained by varying the inclination of the
blastholes.

▪ Toe: In an opencast mine, the Bottom Face of the Bench Crust, is referred to as “Toe”. In other
words, the position formed at the Bottom Face, due to Un-blasted materials, is called Toe.
In places, where the blasting has been done unevenly, due to improper drilling there is a
formation of Toe. Due to this formation, the Shovel experiences problem, in picking up the un-
blasted material from the Bottom face.
In order to remove the problem of Toe, the blastholes of depth
greater than the height of the bench are made in the Initial Benches. Due to this, a bench of less
height is obtained. These kind of drill holes are known as Sub-Grade Drilling. The sub-grade drilling is
around 10% of the length of the holes.

# Precautions to remove the problem of Toe: are as follows:

1. To perform Sub-Grade Drilling
2. To perform Inclined Hole Drilling
3. To properly clean the Holes
4. To keep the depth of holes, 10% more than the height of the benches.
5. To properly drain the water for any water etc.
6. To perform Effective Blasting
7. To use Explosives with High V.O.D. (velocity of detonation)
▪ Statutory Provisions: As per CMR-1957 and MMR-1961 & the DGMS circulars published from time to
time, the following practices are recommended or granted under special permission and
circumstances and at the same time exemption may also be granted by the Regional Inspector in
case of coal mines and Chief Inspector in case of metalliferous mines:
hh. The Regulation 105 says about the Manual Opencast Working, and suggests the following
precautions:
i. In alluvium, soil, morum, gravel, clay or other similar soft ground:
 1. height of the bench shall not exceed 1.5m
2. the breadth (width) shall not be less than the height and,
3. the angle of slope should not exceed 45 degrees from the horizontal.
ii. When a pillar is left in-situ for the purpose of measurement of volume of excavation, its
height shall not exceed 2.5 m and where the height of such a pillar exceeds 1.25m, the
base of the pillar shall not be less than 1.5 m in diameter.
iii. In coal, the sides shall either be kept sloped at an angle of safety not exceeding 45
degrees from horizontal, or the sides shall be kept benched and the height of any bench
shall not exceed 3 metres and the breadth thereof shall not be less than the height.
ii. The regulation 106 says about the Mechanized Opencast Working, and following precaution
have been suggested:
i. In alluvium, soil, morum, clay or other soft ground:
1. Height of the bench shall not exceed 3 meters.
2. The breadth (width) shall not be less than three times the height of the benches.
ii. In coal:
1. The height of the benches shall not be more than the digging height or reach of
the excavation machine in use for digging, excavation or removal
2. The width of the bench shall not be less than:
a. The width of the widest machine plying on the bench plus two metres
b. If dumpers ply on the bench, three times the width of the dumper
c. The height of the bench.

Whichever is more.

jj. In metalliferous mines where ore, ‘float’ and any other mineral are to be worked, the height of
the bench shall exceed 6 m (under special permission maximum up to 7.5 m), the breadth
(width) of the benches shall not be less than 6 m. and the angle of slope should not exceed 60
degrees from the horizontal.
kk. In case of hard and competent overburden, benches a maximum bench height of 7.5 m, a
minimum bench width 15 m (for lower height, lower bench width may also be permitted) and
angle of slope should not be more than 75 degrees from the horizontal. However, in case of a
coal mine a minimum bench width of 1 m may be permitted.
ll. The angle of slope of the spoil bank can be kept maximum up to 37.5 degrees from the
horizontal. The spoil bank face shall not be retained by any artificial means at an angle more
than its natural angle of repose. If any spoil bank exceeds 30 m in height, it shall be benched in
such a manner that the general overall slope of spoil does not exceed 33.7 degrees and height
does not exceed 30 m. The toe of a spoil bank shall not be nearer by less than its height to any
public road, railway, public buildings and other permanent structures, which are not belonging
to the mines owner and a suitable fence shall be erected between the toe of an active spoil bank
and those of public roads, railways or public structures, to prevent any unauthorized person to
approach the spoil bank. Nobody should be permitted to approach the toe of an active spoil
bank where he may be endangered from the material rolling down the face.
mm. Under cutting: No persons should be allowed to undercut any face or side for robbing
coal/mineral which may cause serious overhang. If any undercut is noticed, it must be either
filled up or properly fenced immediately. If it is not possible. nobody should be allowed to take
rest during rest time or take shelter during a heavy rain under the ledges. Strict vigil, proper
supervision keeping the advancement of overburden benches ahead of the mineral benches,
proper fencing, etc. will reduce the accident of side fall occurs due to this reason.
 Unit-1 (LO2)
o Unit Operations Involved: The following unit operations are involved, in the procedure of
performing Opencast Mining in any place:
1. Permissions: In order to establish a mine, permissions have to be taken from the D.G.M.S., prior
to any other operations. After this, the permission of State, Division & District Administration
has to be taken. In case, where the site is located in forest area, the permissions from the
concerned Forest Department has also been taken.
2. Planning: At the planning stage, the following aspects are taken care of:
a. Development of Conceptual Model for mine, and then going for detailed Engineering
Studies
b. Plans, Sections, Reports & Drawings are prepared
c. Planning includes details of construction, development and final exploitation schedules,
manpower, equipment, material energy and budget.
3. Reconnaissance Survey: This survey, is done in the place of proposed, in order to identify any
potential sources of risk to the mine. During this, the area required for office etc. is cleaned
from vegetation etc.
4. Site Preparation: This includes:
a. The removal of vegetation & cleaning the site from any obstruction
b. The topsoil is removed & transported to store it for future reuse or for the direct
replacement
c. In some cases, the initial O.B. could be soft, semi-consolidated or unconsolidated
ground, which could be removed by Dozing, Ripping or Scrapping.
5. Opening up the deposit: Opening up of open pits is done by opening a cut, i.e., Box cut for the
development of first working bench. This cut is first dug in the ground massif, either manually using
conventional tools, or using an excavator & then it is extended & widened and deepened to reach up the
floor of the first bench. In case of rocks, ‘V’, wedge or pyramid cut pattern of holes are drilled & charged
to create an initial excavation. This is known as initial 'BOX CUT’ or "TRENCH”. This trench can be
extended in any direction along or across the longer axis of open pit.

Trenches can be started from outside or inside of overall pit limits. The location of
the box cut will be as per the sequence of mining a deposit. It could be either at its terminal points or at
middle. A trench can serve one bench or several benches or all the benches up to the ultimate pit depth.
When the first bench is sufficiently advanced, the box cut is oriented & extended to the next lower
bench keeping sufficient number of rooms for the top bench & trench approach road to the top bench &
for opening for 2nd bench.

In this way, a no. of working benches is developed & width of box cut should be sufficient
enough to diversity the approach road to all the benches. The box cut, one opening cut, should be
started enough away from the pit limit, so that bottom bench can be reached at the desired slope of the
pit.

6. Development: The development work is begun with putting up the box cut, which gives way to
development of ramps & benches in waste rock, as well as in ore deposit.
Pit geometry will be governed by the geometry of ore body in general & to the dip of the
deposit in particular. Different patterns of ore mining within the pit limits could be single sided, double
sided in longitudinal as well as in the transverse direction, or centralized. In development, the waste
rocks dump yards could be located based on geometry & suitability of the available land in terms of
techno-economical aspects.
7. Production: includes:
a. Drilling & blasting operation
b. Muck handling
 c. Transport system.

All these major operations are further described in detail.

o Box Cut: Opening up of open pits is done by an opening cut for the development of first working bench. The
opening cut is called the box cut. It is one of the two methods of Opening a Mine. In this method, a pit that is
10m. long & 10m. wide, is made at the centre of the place, where the deposit is located.

Box cut (Fig 3.1) is excavated initially down to the floor level of the first bench from
the surface. Thereafter an opening level trench is extended from this opening cut to form the first bench.
The opening trench is narrow keeping due regards of the turning of the machineries used for excavation and
extends along or across the quarriable limit depending on the type of the deposit. When the first bench is
sufficiently advanced, the box cut is oriented and extended to the next lower bench keeping due regard of
the sufficient number of rooms for the approach road to the top (1st) bench and for opening trench for the
2nd bench.

This way a number of working benches are developed and the width of the box cut
should be sufficient enough to diversify the approach road to all the benches. If number of benches are
developed from one opening cut, the cut should be started enough away from the pit limit so that bottom
bench can be reached at the desired slope of the pit. This type of opening cut may be very long and may be
curved depending upon the shape and extent of the deposit. For opening up in hilly deposit, a central trench
cut is given across the top level for the first bench or from one side in the same contour level forming a
length of face which will give the required production rate.

# Location of Box Cut: The location of the box cut depends upon the following factors:

1. Minimum cost of haulage within the open-pit and outside the pit to the desired place e.g., preparation
plant, sidings etc. or mineral and overburden dump. However, if the place of mineral siding and
overburden dump is far apart. separate opening cuts one for mineral and the other one for overburden
is to be formed.

2. If shifting of box cut is not necessary, it is to be located at the boundary of the mineral deposit/property.
In dipping deposit, this boundary should be the mineral outcrops or where ratio of overburden: ore is
least, location of the box cut should be in the middle of the boundary of the reserve for minimizing the
haulage cost.
3. In horizontal deposit, the box cut can be located in the middle of any boundary depending on the
location of the place of destination of the overburden and mineral disposal. The opening trench in
continuation of the box cut offers two working faces of adequately long for the desired production on
each level. The site of the box cut should be stable and free from geological disturbances.

4. The site should be at the rise side to guard against the flooding of the mine.

5. The site should be selected where the construction of approach road is very convenient.

6. The site should be selected preferably where the deposit has high grade mineral to compensate the
development cost. Location of the box cut should be such that it serves the purpose of maximum mining
area.

However, if the mining area is extensive, the total

mining area to be divided into sub-areas and they are to be opened up separately. Due to the wider and
deeper reach draglines are the most suitable equipment for making a box cut. Standing on the surface they
can load materials from the box cut and unload to the railway wagon.

# Types of Box Cut: There are two types of the box cuts:

1. Internal Box Cut: If the box cut is placed vertically above the deposit, then it is called Internal Box Cut.
2. External Box Cut: If the box cut is not located, vertically above the deposit, but is situated within the pit
limit, it is called External Box Cut.

Generally, the box cuts are made in the waste material, as the e deposit is at
some depth from the surface. But in case, of a deposit, that has an outcrop on the surface, the Box Cut is
made is directly into the Deposit/Orebody. In order to access the deposit, multiple box cuts may be required
to be made.

# Dimensions of the Box Cut: are as follows:

1. The Length of the Opening Cur: is equals to the breadth of the cut
L=B
2. The volume of the material extracted from the Cut:

𝑐𝑜𝑠ℎ2
𝑉=
𝐼
 ℎ
(𝑉 + ) 𝑚3
𝑡𝑎𝑛𝛼

# Access Trench Method: It is a method of opening up a deposit, in an opencast mine. This method is not
generally used, due to the reasons that it makes the mining conditions difficult, and it is costly to make the
roads, using this method.

In this method, a tunnel or incline, is made at one corner of the mining place, and without disturbing
the floor of the incline, a bench is made from the direction of the roof. The bench is made in the shape of
“U”. Then, the incline is deepened, and a second bench is made. In this way, the successive benches are
made.

The speciality of this method, is that all the

benches are connected to the same road, which travels inwards. The gradient of this road is not kept less
than 1 in 4.

o Opencast Layout:
# Figure: on next page:

# Factors Determining Choice of Layout: are as follows:

1. Topography: The topographical conditions if the area to be mined out, is the major factor that affects the
choice of the layout
2. Geology: The geological conditions, and the presence of various geological disturbances etc. also affects the
choice of the layout.
3. Extent of the Deposit: The geographical extent and the spread of the deposit plays a key role in the process
of the selection of the layout.
4. Type of the Deposit: The type of deposit (bedded, massive, vein etc.) is one of the key factors that is
determined, during the selection of the layout.
5. Grade of the ore: The grade of the ore, present in the deposit is also considered importantly, while
performing the selection of the layout.
6. Production Rates: The rates of production of the ore, in the particular area, is also considered in the
selection of the layout.
7. Mining Costs: The costs incurred, in order to mine out the deposits are a key consideration, while the
selection of the layout is carried on.
8. Processing Costs: The rates of processing of the ore, in the particular area, is also considered in the selection
of the layout.
9. Metal Recovery: The rate of recovery of the metal from the ore is also taken care of, while performing the
selection of the layout.
10. Bench Height: It is a major consideration in the selection of layout. It is considered in order to determine the
stability of the layout.
11. Pit Slope: It is also a major consideration in the selection of layout. It is considered in order to determine the
stability of the layout.
12. Stripping Ratio: In order to detect the economic feasibility of the layout, the Stripping Ratio of the area o&
the ore, is considered, while selecting the layout.
13. Marketing Consideration: The availability of the markets for the particular area, and their conditions are
also determined, while performing the selection of the layout.
14. Cut Off Grade
15. Localization of the deposit
16. Property Boundaries

# Precautions: the following precautions are kept in mind:

1. The galleries of underground mines, located near an opencast seam, shall be closed with the help of an
Explosion Proof Stopping.
2. Stone Dusting is done in such galleries

Geography
Geology
Extent
Type
Grade
Cut-off grade
Metal recovery
Production cost
Mining cost
Processing cost
Marketing consideration
Bench height
Pit slope
Stripping ratio
Boundaries
Localization
3. For blasting in such areas, the blast holes made in the overburden and the benches of coal, shall be made
such, that are not located over the Galleries.
4. Pilot Holes are made on the Overburden bench, located over a coal bench, to know the thickness of the
parting left between the Coal Bench and the Underground Seam.
5. The blast hole in the first bench over the coal, is filled with wet sand, up to a height of 0.6m. and is then
charged with explosive.
6. As far as possible, the galleries of the UG seam are filled with Overburden, before introducing heavy
machines on the coal bench.
7. As far as possible, the shovel shall be not placed directly over the gallery.

# Machines: are as follows:

1. Coal Bench: for the coal benches, the machines are as follows:
a. Shovel - 1
i. Size – 4-6m3
ii. Boom Height – 10m.
b. Dumpers - 6
i. Size – 35 tons
c. Drill Machine - 1
i. Size – 200mm.
d. Dozer - 1
e. Pay Loader – 1
2. Overburden: The machines for the overburden, are as follows:
a. Shovel - 3
i. Size – 4-6m3
ii. Boom Height – 10m.
b. Dumpers - 18
i. Size – 35 tons
c. Drill Machines - 3
i. Size – 250mm.
d. Dozers - 3
e. Scrapers - 2
i. Size – 11m3
3. Other Equipment: are as follows:
a. Grader
b. Road Roller
c. Crane
d. Explosives Van
e. Pump
f. Water Tank

# Manpower: is as follows:

1. Per Shift

S. No. Designation No. per shift

1 Overman 1
2 Mining Sirdar 1
3 Shovel Operator 4
4 Shovel Helper 4
5 Dumper Operator 24
6 Drill Machine Operator 4
7 Drill Machine Helper 8
8 Water Tanker 4
9 Scraper 2
 10 Grader 2
11 Road Roller 2
12 Pay Loader 2
13 Dozer 6
14 Pump Man 4
15 Pump Helper 2
16 Electrician + Helper 3
17 Mechanical Foreman 1
18 Mechanical Fitter + Helper (Dumper) 6
19 Mechanical Fitter + Helper (Shovel) 6
20 Other Staff 10
Total 96

2. General Shift

S. No. Designation Number

1 Blasting Overman 4
2 Blasting Gang 20
3 Maintenance Shovel 6
4 Maintenance Dumper 10
5 Foreman In-charge 1
6 Electrical Supervisor 1
7 Other Works 10
Total 52

Total Manpower in all shifts: 3 * 96 = 288

Manpower in General Shift = 52

Total = 340

𝑂𝑢𝑡𝑝𝑢𝑡 𝑃𝑒𝑟 𝐷𝑎𝑦

Output Per-man Per-shift (O.M.S.) = 𝑁𝑜.𝑜𝑓 𝑝𝑒𝑟𝑠𝑜𝑛𝑠 𝑒𝑚𝑝𝑙𝑜𝑦𝑒𝑑 𝑝𝑒𝑟 𝑑𝑎𝑦

3400
= 340

= 10 tons per-man per-shift

o Overburden Excavation: The removal of overburden is a very important operation in surface mining system.
The method of excavation of overburden depends mainly upon the following main factor:
1. Thickness, dip and depth of the overburden.
2. manner of occurrence of the deposit, (i.e., whether the deposit have occurred under the flat surface
terrain or over the hilly terrain)
3. the surface topography,
4. the environment conditions
5. the ground conditions,
6. the production requirements,
7. the geo-technical parameters of rock like:
 a. the compressive strength,
b. the shear strength,
c. the tensile strength,
d. the modulus of elasticity,
e. mineralogy,
f. lamination,
g. massiveness,
h. toughness
8. geological disturbances,
9. Watery condition
10. the stability of the overburden and the spoil benches, etc.

The selection of excavating equipment like dragline, shovels, front end load bucket
wheel excavator, scraper along with ripper, surface continuous miner, etc. also depend upon the above geo-
technical parameters of rocks. The physical characteristics like thickness, floor conditions of geological
structures are also considered.

The Excavated overburden are disposed-off either within the void

created by the worked-out areas, or away from the open pit. The number of entries (ramps, etc.) to enter in
or exit from the mine depends upon the following factors:

1. length of the mine on the strike direction

2. amount of required production
3. the requirement for separate route for mineral and overburden transport
4. Distance of Haul
5. System of handling of overburden, soil, rocks and also of minerals.
6. Working of number of faces simultaneously
7. Overall economy.

The location of the entries is depended upon the following factors:

1. Surface Topography
2. Shape of the mine
3. Length of the mine along the strike
4. System of waste disposal & location of disposing site
5. The required gradient of the haulage system.
6. Depth of the working
7. Requirement of the production
8. Overall economy.

o Disposal Of Overburden: The process of the disposal of the overburden can be carried out, inn either of the
three methods:
1. Disposal of the overburden rocks away from the Pit: This type of practice is generally adopted under
the following conditions;
a. The deposits are thick and steeply inclined
b. Deposits occur in hilly terrain
c. Ponds, lakes, any unproductive low-lying topography area to be filled up for special use or
cultivation
d. For making embankment
e. Extensive removal of overburden is desire to obtain huge amount of productions
f. In the deep pits
 2. Disposal of Overburden from the deposits occurring in the Hill Top: The disposal of overburden is easily
done by discharging the waste rock from the hill top. The overburden rocks are taken to a particular side,
running all along the face length either simply by a truck road, or by a railway line, and are dumped to fill
up the valley area for reclamation.
The disposal of overburden directly in the downhills side will cause
degradation of mining environment like soil erosion, water pollution, ecological imbalance, destruction
of trees, fertile lands, hills, slope lands etc. The pushing of the spoils in the downhill side should be in the
steeper portion and shall be in a controlled manner to fill up the valleys near to the mine away from the
cascade, seepage or drainage oof water from other sources.
Before filling the valley, the quantity of swelled overburden volume is to be determined, to
fill up the valley as well as disposing it off outside the extracted area. The valley is to be prepared first, to
accept the dump material. Then, the top soil forms the hill top area, is to be stripped out and
transported and stacked in a separate place away from the valley. After making a trench in a suitable
place, the overburden material is dumped over the prepared area.
3. Disposal of Overburden Material by Vertical Inclined Shafts and Drifts: In steep hilly terrains, there are
several restrictions on the environmental pollutions like vibration and noise form the plying of heavy
trucks and vehicles, production of noise form the running of mining equipment, air, water and soil
pollution etc.
However, where a heavy amount of production is
necessary, a vertical or inclined shaft around 8-10m. in diameter, maybe sunk through the ore body, at
an optimally located place. The shaft ends at the foot hill of the mountain, where a tunnel drifted from
the foot of the mountain may be connected to the bottom mouth of the shaft fitted to the chute.
A primary crusher is installed near the top mouth of the shaft, to crush the rock or
make a particular size fragment from it, and discharge it into the shaft. The crushed ore or rocks are
thereafter transported by a belt-conveyor installed in the tunnel.

o Overcasting: The method of Over-Casting has the following variations:

1. Over or Side Casting by Dragline: The dragline is the most common equipment used for the Over-
Casting/Side Casting process, because of their greater reach, greater flexibility, being unaffected by bank
slides, water runoff and seepages etc.
Also, if required, the boom can also be extended for larger reach. The height and width of
the bench, should be optimum to match the capacity of the dragline and its dumping radius. A stripping
ratio around 2.5:1 (overburden: mineral) is very good for reclaiming the excavated areas.
The side casting system adopted by the
draglines, can be widely grouped as:
a. Simple: In this, the Dragline is installed over spoil heap after levelling the same. Dragline
excavates overburden rock over the coal seam (portion A) and dump it over the spoil heap
(portion B) for exposing the coal seam. The exposed coal will be extracted by separate benches
with the shovel dumper combination.
b. Extended Bench: In this, the dragline is located over the levelled spoil heap and excavates the
overburden rock over the coal seam and dump the same over the spoil dump for exposing coal
seam. The exposed coal will be worked out separately with the shovel dumper combination. In
this system dragline increases its reach at the cost of its productivity. While taking thee soil/rock
from the top; the efficiency of the draglines decreases by 70%.
c. Horse Shoe
d. Tandem: In this system, two Draglines are used. Dragline No. 1 is located over the surface which
excavates overburden rock over the coal seam and dump it over spoil heap (Zone 1). The
dragline no. 2 is located over the levelled spoil which excavates the remaining overburden left
over the coal seam and dump it on the spoil (Zone2) and also it rehandles the spoil of Zone1 to
dump it in proper place. The exposed coal will be worked out separately with the shovel dumper
combination.
2. Over Casting by Shovels: It is not very common to overcast the overburden by means of shovels, as both
its digging and dumping radius are short. However, sometimes, in a single overburden bench, blasted
muck maybe loaded and over casted by a single or double pass system, after moving the shovel to a
greater amount at the face to cover entire, up to the face width.
Special precautions are to be taken, during over casting by means of a shovel, which are as
follows:
a. The distance between the bottom of the mineral or coal and the bottom of the dump depends
upon the mode of the method of transport system, the mode manner and the size of the
fragmentation, production required, bench parameters.
 b. In case, there is a two-way truck haulage and the mineral is excavated directly by means of a
shovel without drilling and blasting operations.
c. In case minerals are to be blasted, the fragmented particles will be scattered at least inn a radius
of twice the width of the blasting face.
3. Over Casting by Bucket Wheel Excavator: The Bucket Wheel Excavator along with the combination of
spreader can effectively be used for over casting the loose, soft and the thick overburden. However, the
hard rocks should be pre-blasted with the help of light charge for fragmentation. A B.W.E. gives
continuous output and has low operating and power costs. They have excellent loading and dumping
characteristics & facilities.
The continuous Bucket Wheel Excavator systems may have the following
combinations:
a. Bucket Wheel Excavator – Conveyor Bridge
b. Bucket Wheel Excavator – Intermediate Mobile Belt Conveyor
c. Cross Pit System (BWE-Intermediate Mobile Belt Conveyor-Stacker)
d. Bucket Wheel Excavator – Double Line Locomotive System.

o Layout Problems:
1. Limestone
a. Question: Draw a layout of a surface limestone mine with an output of 1000 tonne/day in a flat
deposit. Assume your own conditions.
b. Introduction: Let the following points be considered:
i. The thickness of deposit: 30m.
ii. Gradient: Flat
iii. Overburden: Top Soil with thickness around 3-4m
iv. Deposit is quite sufficient and the life of the mine is around 30yrs.
v. Railway station is very nearby

The cement plant is located within 2km. distance. Its life is 30yrs. If the cement plant is not located
very near to the limestone mine, mechanization of mine will not be possible since cost of limestone
is not very much.

c. Description of the Mining Layout: for the production of the ore of 10000tonnes/day:
i. Number of production benches/faces required:
ii. Height of the mineral bench: 10m.
iii. Slope of the mineral bench: 65 degrees
iv. Transporting Distance: 200m.
v. Cost of Limestone per tonne= Rs. 109
vi. Total Shifts: 4
1. Production shifts: 3
2. Maintenance Shift: 1
vii. Shovel Capacity: 3.2m3 – 3.6m3
viii. Maximum amount of investment for plant 7 machinery in 1st phase (10yrs.) = Rs. 39.24 crore.
d. Figure: is as follows:
e. Main Equipment: as follows:

f. Other Equipment to be procured: are as follows:

i. Dozer cum Ripper Combination – 2
ii. Compressor (7-9m3/min. capacity & number depends upon the quantity of compressed air
required)
iii. Jack Hammer Drills - 3
iv. Water Sprinkler – 1
v. Portable Workshop containing all the equipment for repairing, including portable gas cutting
arrangements, electric welding set etc.
vi. Electrical Equipment in the substation
vii. Exploders, circuit testers
viii. Safety Appliances & equipment
ix. First Aid & medical Appliances
x. Office Equipment
g. Method of Working: First of all, remove any tree or bushes present over the site of the quarry, grub
the tree stumps either with the help of bulldozer or simply by drilling and blasting operation. Scarify
the ground. After scarification is over doze the loosen rock with the help of bulldozer. Make clean
the mining site. A box cut is to be excavated over the soil and the limestone deposit along the
longitudinal direction dividing equally the whole property. Box cut is to be made by drilling and
blasting method. Mucking and transporting of both overburden top soil and the limestone will be
done with the help shovel and dumper combination along with the dozer. When sufficient bench
width will be available, start winning of limestone mineral and at the same time start excavation of
dozing and stacking properly the top soil around the boundary of the deposit.
This dozing is to be done in
the transverse direction in both the faces of the bench. A clearance space of around 37m to 40 m
from the edge of the bench is necessary. Rail tracks are laid longitudinally as shown in the figure 4.1
leaving a clear space around 3 to 15 m from the edge of the bench. This is done taking due
consideration of scattering of blast resulted muck pile so that the later does not affect the rail lines
 and hinder the movement of railway wagons. Shunting scope is provided to divert empty wagons to
other face or to store empty wagons there itself. Position of shovel for loading ore onto wagons,
movements of face in the longitudinal direction and movement of bench in the transverse direction
has been shown in the figure 4.1. After taking one slice of 12 m width of bench all along the
longitudinal length rail lines are to be transversely shifted keeping a minimum distance of 13m to 15
m away from the bench edge. After the finish of first bench. second and third benches are to be
worked out successively. The ore to be transported by surface locomotives and wagons, which are to
be tippled by a tippler inside the cement factory.

h. Production & Productivity:

i. Production (given) = 10000t/day
ii. Expected Overall Productivity (OMS) = 30 tonnes/man/shift

It is known that the overall OMS of an Opencast Mines =

𝑐𝑜𝑎𝑙 / 𝑚𝑖𝑛𝑒𝑟𝑎𝑙 𝑝𝑟𝑜𝑑𝑢𝑐𝑡𝑖𝑜𝑛 𝑖𝑛 𝑡𝑜𝑛𝑛𝑒𝑠 +𝑜𝑣𝑒𝑟𝑏𝑢𝑟𝑑𝑒𝑛 𝑖𝑛 𝑚3
𝑚3
………(1)
𝑇𝑜𝑡𝑎𝑙 𝑀𝑎𝑛𝑠ℎ𝑖𝑓𝑡 (1+𝑑 × 𝑠𝑡𝑟𝑖𝑝𝑝𝑖𝑛𝑔 𝑟𝑎𝑡𝑖𝑜 𝑖𝑛 )
𝑡

Where, d = density of the Limestone

Let, Ultimate stripping ratio be: Cubic meter of Overburden: Tonne of Mineral = 0.11: 1

Let, Density of Limestone = 2.6t/m3

Putting all the above value in the equation 1

10000 + 2.6 × 1350
30 =
𝑇𝑜𝑡𝑎𝑙 𝑚𝑎𝑛𝑠ℎ𝑖𝑓𝑡 (1 + 2.6 × 0.11)
10000 + 3510
𝑜𝑟 𝑇𝑜𝑡𝑎𝑙 𝑀𝑎𝑛 𝑆ℎ𝑖𝑓𝑡 = = 350.18 ~350 𝑝𝑒𝑟𝑠𝑜𝑛𝑠
30 (1 + 2.6 × 0.11)
Hence, estimated minimum manpower of the mine = 350 persons

2. Copper:
a. Question: Draw a layout for a copper mine with an output of 5000 Tonne per day in a flat deposit.
The thickness of the copper deposit is 10m. & the thickness of overburden is 10m. Assume your own
conditions.
b. Introduction:
i. Output Per Day = 5000 tonnes
ii. The thickness of the copper deposit: 10m.
iii. The thickness of the overburden: 10m.
iv. Life of the mine- 10yrs.
v. Railway Line Distance- 2 km.

Let following are considered:

i. Deposit is quite sufficient and the life of the mine will last around 10 yrs.
ii. Railway lines are 2km. away

But due to the nature of the surface terrain outside the mine, rail connection up to the mine is not
possible. Hence it is decided that a dump yard along with small ore handling plant and bunker is to
be built up near to that point (i.e., near too the railway line).
c. Description:
i. Production of Ore= 5000 tonnes/day
ii. Height of the bench= 10m.
iii. Slope of the bench= 65 degrees
iv. Height of the Overburden= 10m.
v. Width of the Overburden= 30m.
vi. Slope of the Overburden= 65 degrees
vii. Cost of Copper Ore= Rs. 650 per tonne
viii. Maximum Investment to be dine for purchasing plant and machinery= around Rs. 110.5
crore
ix. Total Shifts: 2
x. Dumper Capacity: 35 tonnes
xi. Shovel Capacity: 4.6m3

d. Main Equipment:
are as follows:

e. Other Equipment Required: are as follows:

i. Primary Crusher System
ii. Ore Handling Plant including small bunker for loading of ore over the railway wagon.
iii. Compressors (7-9m3/minute capacity & Number depending upon the amount of
compressed air required)
iv. Jack Hammer Drills: 3
v. Portable Workshop containing equipment required, including portable electric welding set,
gas cutting arrangements etc.
vi. Electrical Equipment in the sub-station
vii. Safety Appliances & Equipment
viii. First Aid & Medical Equipment
ix. Office Equipment-
x. Mobile Crane
xi. Welding Truck
xii. Giraffe Truck
xiii. Cable Reel Truck
xiv. Tyre handler
xv. Power Sweeper
xvi. Steam Cleaning Machine
xvii. Wheel Motor Dolly

f. Method of Working: First of all, remove any trees or bushes present over the site of the quarry, grub
the tree stumps either with help of bulldozers or simply by drilling & blasting operations. Scarify the
ground. After scarification is over, doze the loosen soil/rock with the help of bulldozer. Make clean
the mining site. A box cut is to be excavated over the overburden rock (granite) and chalcopyrite
mineral body along the longitudinal direction dividing equally the whole property by the drilling and
 blasting method. Mucking and transporting of both the overburden rock and the ore will be done
with the help of shovel and dumper combination. When sufficient bench width will be available,
start winning of overburden in full swing first in one direction and make a clearance of nearly 30m to
start the operation of the winning of chalcopyrite mineral. For proper utilization of the 4.6 m3 bucket
capacity rope shovel, drilling and blasting operation shall be carried out in such a manner and in such
a width that a 15m to 16m width of the face is available for excavation by a shovel. In the figure 4.2
it is clearly shown that a ramp (gradient 1 in 10) around 6 to 7 time the maximum width of the
dumper is excavated to reach the overburden bench and similarly a ramp (gradient 1 in 10) around 4
to 5 times the maximum width of the dumper is made from the floor of the overburden bench to the
floor of the mineral bench. In the first phase after advancing face no. 1 of both the mineral and
overburden bench to a considerable distance, start winning of face no. 2 of both the overburden and
mineral bench to make them stand-by and for better quality control. Select proper dumping place of
the overburden only in one side (right hand side, i.e., in the side of the face-1) outside the boundary
of the mine. But when sufficient space will be available on the floor of the mineral bench after
extraction of minerals, start dumping of the overburden over the floor of the mineral bench starting
from the middle point. It will minimize the hauling distance of the overburden as well as will
eliminate rehandling of that overburden. Try to finish both the overburden and mineral bench faces
(no.1) in one direction only which will systematize the reclamation of overburden rock in the
opencast mine. Overburden rocks of face no. 2 can directly be dumped on the floor of the mineral
bench.
Ultimate slope of the pit will be around 40 to 45°. Fill up the dematerialized
area with the overburden rock stacked outside the boundary of the mine after extraction of the total
mineral. Movement of the faces is in the longitudinal direction and the movement of the benches is
in the transverse direction. The ore is transported by the dumpers and dump over the primary
crushing plant to crush the ore in the size of 150mm. From the crusher plant, ore is transported by
the belt conveyor to the material handling plant. From the material handling plant ore is loaded onto
the railway wagon through the bunker by apron feeder. Floor of the overburden bench shall be
inclined at an inclination of 1 in 20 towards the extraction side. During excavation, the floor of the
mineral shall be inclined towards the sump so that water during the monsoon season can directly fall
over the sump. If necessary, right from the surface up to the sump, drains and ditches are to be
formed for free flowing of water.

g. Production & Productivity:

i. Production = 5000 tonnes per day
ii. Overall Productivity (OMS) = 25 tonnes/man/shift
iii. Minimum manpower required = around 200 persons

3. Coal:
a. Question: Draw a layout for a coalmine with an output of 1000 Tonne per day in a deposit having a
gradient of 4 degrees. Assume your own conditions.
b. Introduction: following points are considered
i. Output Per Day = 1000 tonnes
ii. The thickness of the coal deposit: 10m.
iii. Dip of the deposit is 4 degrees and the deposit is outcropped
iv. The average thickness of the overburden: 3-4m.
v. Life of the mine- 12yrs.
vi. Railway Line Distance- 1.5 km.
c. Description:
i. Production of Ore= 1000 tonnes/day
ii. Height of the bench= 10m.
iii. Height of the Overburden Bench= 3-4m.
iv. Width of the Production & Overburden Bench= 10-15m.
 v. Cost of Coal= Rs. 1000 per tonne
vi. Maximum Investment to be done for purchasing plant and machinery= around Rs. 24 crores
vii. Total Shifts: 4
1. Production- 3
2. Maintenance- 1
viii. Height & Slope angle of the coal benches are kept as follows:

Bench Height Slope Angle

1st Bench 3.0m. 50 degrees
2nd Bench 3.5m. 50 degrees
3rd Bench 3.5m. 50 degrees

Since the deposit is outcropped, gradient is mild and production is not very much, manual method of
working will be the best and cheapest proposition. Direct Rope Haulage will be the cheapest mean of
transport system right from the pit to the tippler near to the railway siding located 1.5km. away
from the mine.

d. Main Equipment: are as follows:

e. Other Equipment Required: are as follows:

i. Compressors (7-9m3/minute capacity & Number depending upon the amount of
compressed air required)
ii. Jack Hammer Drills: 5
iii. Portable Workshop containing equipment required, including portable electric welding set,
gas cutting arrangements etc.
iv. Electrical Equipment in the sub-station
v. First Aid & Medical Equipment
vi. Office Equipment
vii. Tippler
viii. Exploder, Circuit Testers etc.
f. Method of Working: First of all, remove any trees or bushes present over the site of the quarry, grub
the tree stumps either with the help of bulldozer or simply by drilling and blasting operation. Scarify
the ground. After scarification is over, dove the loosen soil/rock with the help of bulldozer. Make
clean the mining site. A box cut is to be made over the file hard top soil and coal outcrop. The box
cut is made along the strike direction starting from the outcrop of the coal deposit and to be
touched up to the ends of the boundary of the mine. Haulage machine including the haulage
building, main haulage tracks, face haulage track, shunting facilities, etc. are to be constructed very
carefully. It would be better if main haulage tracks are laid at one end of the mine. Box cut can be
made by drilling and blasting operation or simply by manual excavation. After cut making is over
(making trenches in the strike-direction), winning operation is to be carried out by Simply drilling and
blasting operation and loading is to be done manually over the 1m3 mine-tubs. Jack hammer drills
can be employed in the first phase of operation like making of box cut and for advancing the face to
a considerable distance. One jack hammer can drill on an average 50 number of holes each of 1.5m
length. Spacing of holes are 1.2m. to 1.5m. and burden of holes are 1m. to 1.2m. High strength
special Gelatine explosive of amount 200 to 250 gm/hole is generally used A powder factor of 10 to
11 Tonne/kg of explosive easily can be obtained. A team of 2 miners are allotted to a space of nearly
2m along whole length of the face in the strike direction to load coal over the 1m. mine tubs. But at
the initial starting point of box making operation, coal is to loaded over the tipping trucks of 8 Tonne
capacity.

As the extraction of coal goes on, a time will come when the height of the bench will be more than
3m. At that time a 2nd coal bench will be formed and when this bench will also exceed 3.5m., a third
coal bench will be formed. During the progress of the mining activity overburden top soil will also be
touched and an overburden bench has to be formed (Fig 4.3). This overburden top soil will be
transported with the help of same haulage system and be dumped over the excavated de-coaled
area systematically for the reclamation of land (Fig 4.3). Since there is no tippler over the reclaimed
land, side tipping mine tub should be used in this case. All the bench floor shall be inclined at an
inclination of 2 degrees towards the excavation for maintaining a good drainage system. The haulage
tracks are laid as shown in the fig 4.3 and the minimum inclination of the main haulage track is
maintained at a gradient of 1 in 10. The level haulage tracks near to the faces are kept in a sufficient
distance apart to avoid any damage. Proper shunting facilities are provided near to the boundary
(outcropped) side of the mine and also, if necessary, a small haulage may be installed there. Blasting
is done in the longitudinal direction in single row in a slice ranging in between 2.5m to 3 m. After
taking 1 to 2 slices the tracks of the mine cars are to be shifted towards the face to minimize the lead
distance. The trains of mine tubs (around 20 to 25 in number) are hauled by the direct rope haulages
placed at a distance around 1.5 Km away. The loaded train when reaches near to the unloading
point, is then shunted and lowered down a little bit and in place of it, empty train is connected with
the haulage rope and lowered down to the faces of the benches for reloading purpose. The loaded
tubs are tippled by the Tippler to unload over a bunker from where coal is transported by the railway
wagons.

g. Production & Productivity:

i. Production (given) = 1000t/day
ii. Expected Overall Productivity (OMS) = 3 tonnes/man/shift

It is known that the overall OMS of an Opencast Mines =

𝑐𝑜𝑎𝑙 / 𝑚𝑖𝑛𝑒𝑟𝑎𝑙 𝑝𝑟𝑜𝑑𝑢𝑐𝑡𝑖𝑜𝑛 𝑖𝑛 𝑡𝑜𝑛𝑛𝑒𝑠 +𝑜𝑣𝑒𝑟𝑏𝑢𝑟𝑑𝑒𝑛 𝑖𝑛 𝑚3
𝑚3
………(1)
𝑇𝑜𝑡𝑎𝑙 𝑀𝑎𝑛𝑠ℎ𝑖𝑓𝑡 (1+𝑑 × 𝑠𝑡𝑟𝑖𝑝𝑝𝑖𝑛𝑔 𝑟𝑎𝑡𝑖𝑜 𝑖𝑛 )
𝑡

Where, d = density of the Limestone

Let, Ultimate stripping ratio be: Cubic meter of Overburden: Tonne of Mineral = 1.2: 1

Let, Density of Limestone = 1.3 t/m3

 Putting all the above value in the equation 1
1000 + 1.3 × 1200
3=
𝑇𝑜𝑡𝑎𝑙 𝑚𝑎𝑛𝑠ℎ𝑖𝑓𝑡 (1 + 1.3 × 1.2)
1000 ÷ 1560
𝑜𝑟 𝑇𝑜𝑡𝑎𝑙 𝑀𝑎𝑛 𝑆ℎ𝑖𝑓𝑡 = = 333.33 ~335 𝑝𝑒𝑟𝑠𝑜𝑛𝑠
3 (1 + 1.3 × 1.2)
Hence, estimated minimum manpower of the mine = 335 persons

4. Iron Ore Deposit:

a. Question: An iron ore deposit of 10m. thick and dipping at angle of 45 degrees and outcrops on the
top of the hill. Overlain soil cover is 5m. thick. Give Layout of a quarry and requirement of machinery
to produce 1000 tonnes per day of saleable iron ore. Assume any condition if necessary.
b. Introduction: The following conditions are considered:
i. Thickness of Iron Ore Deposit: 10m.
ii. The dip of the Deposit: 45 degrees
iii. Overburden (Soil) Thickness= 5m
iv. Deposit is outcropped on the top of the hill but the deposit ends at a point 10m. above from
the foot of the hillock
v. Deposit is quite sufficient and the life of the mine is expected to be around 10 yrs.
vi. Railway Link is available 2km. away from the hillock
c. Description of the Layout: are as follows:
i. Selling Price of Ore: Rs. 350 per tonne
ii. Production: 1000 tonnes/day
iii. Minimum Investment required to be made for plant and machineries= Rs. 8.4 crores
iv. Density of Ore= 5 tonne/m3
v. Thickness of the Overburden= 5m.
vi. Width of Overburden= 7m.
vii. Width of the Mineral Bench: 21m.
viii. Height of the Overburden Face: 3m.
ix. Slope of Mineral & Overburden Bench= 45 degrees
d. Method Of Working: First of all, remove all the trees and bushes from the mining area, as well as
from the slope (mining side) of the hillock. Grub the tree stumps manually with the help off picks or
by blasting and remove them. Clean the mining site. Make the box cut by excavating trench over the
soil along the strike direction. Make flume bunker and loading arrangements. Barricade the railway
tracks, so that the overburden soil cannot fall over them.
After making trench, excavate the full width of the overburden soil manually
by picks and arrange to throw it at the foot of the hillock manually by the hand shovel and baskets
by two teams in each shift comprising of two to three miners per team. When the overburden face
will move sufficiently in the strike direction, start extraction of the iron ore by drilling and blasting
method up to the full width of the ore body to some distance, so that tracks for the mine tubs can be
properly laid over that void space as shown in figure. With the help of jack hammer, start the drilling
operation for blasting.
 e. Main Equipment: are as follows:

f. Production & Productivity: is as follows:

i. Production (given)= 1000 tonne/day
ii. Expected Face OMS= 10 tonne/man/shift
iii. Expected Overall OMS= 6 tonne/man/shift
iv. Estimated Requirement of Total Manpower in the mine= 170 persons

Unit 2 (LO2)

o Classification of Mine Machinery: Just like the case of an underground mine, the machines used in a surface
mine are also categorised in various categories, which are as follows:
▪ Production Machines: The machines, that are used to directly provide us with the production, are
referred to as “Production Machines”. Just like there are machines like Shearer and Continuous
Miners, which are used for the purpose, there are machines like Surface Miners, Draglines, Bucket
Wheel Excavators, that are kept in this category.
▪ Loading Machines: The machines that are employed with the sole purpose of loading the broken
ore/overburden in the haulage units, like tubs, are referred to as “Loading Machines”. The machines,
like LHD & SDL, used underground, are included in this category, Similarly, the machines like Shovel,
Stage Loader, Pay Loader etc. are used for the same purpose, and are kept in the same category.
▪ Transportation Machines: The machines that are used for the purpose of the transportation of the
material, in different parts of the mine, are called “Transportation Machines”. The machines, like
Conveyors, Rope Haulage or Locomotives are used as Transportation Machines. Similarly, the
machines like Conveyors, Trucks, Dumpers, Tipplers etc. are used for the same purpose in a surface
mine, and are kept in the same category.\
 ▪ Other Machines: The machines, that are used for all other purposes, other than the ones listed
above, are kept under this category. Machines like Exploder, Drill Machines, etc. that are used both
at surface and in Underground mines, are categorised in it.
o Selection Of Machinery: In every mine, the criterion for the selection of the different categories of
machines, are different, However, the general factors that are considered for all machines, are as follows:
▪ Production: The first and one of the most important criterions for the selection of any machine, is
the production required. For large production, Large Size machines are used and vice versa.
▪ Nature of Work: Another important factor that is considered. If the material is to be loaded at the
same place, where it is lifted off, Shovels are used. But in case, where it needs to be transported to
some distance, before being loaded, Payloaders are used.
▪ Next/Previous Connecting Machine: In the selection of a machine, the specifications of the
machine, that is just preceding it, is also considered. If it is not done, it greatly reduces the
productivity, either by wastage or by leaving empty space in the machine.
▪ Fragmentation: The size of the material, that has to be broken, is also considered. If the size of the
material that needs to be broken, is small, small size machines can be used. Vice Versa.
▪ Reserve/Deposit: The amount of Reserve/Deposit available at the site for working, is also
considered during the selection of the machines. In case of a large deposit, small size machines are
not used and vice versa.
▪ Cost: The cost of the machine is also considered. It is seen that whether the cost of the machine is
recoverable in the future or not.
▪ Stripping Ratio: If the amount of Overburden that is required to be removed, is greater, the large
size machines are used.
▪ Ground Conditions: The various parameters related to Ground Conditions, like the geological
disturbances, geotechnical information, etc. are also considered.
o Some Other Factors related to Selection of Excavators: are as follows:
▪ Method of Stripping & Disposal of Overburden
▪ Horizontal & Vertical Reach & Reach below the ground.
▪ Operating Conditions and Pit Geometry
▪ Manoeuvrability
▪ Tonnage Required (capacity) to be handled per unit time
▪ Load Bearing Capacity of the floor rock
▪ Power Supply system and arrangements
▪ Surface Topography
▪ Flexibility of Operations
▪ Cost of the machine and availability of the capital & spare parts
▪ Ease of Maintenance
▪ Statistics and Machine Relatability
▪ Operating and Maintenance Costs
▪ Secondary use of Excavator
▪ Climatic Conditions
▪ Overall Safety

o Shovel: A Shovel is an equipment which excavates the rock or ore by digging from its operating base to
upwards (stripping shovel in this case) and dump it in either on a dumper or railway wagon or over the spoil
dump for back-filling after swing itself within its limit. It is a highly productive machine and capable to handle
all types of ores, rocks ranging from fine to very hard blocky lumps, has lower operating costs, higher
production and productivity etc. It requires lower power and has less wire rope cost. It also requires less
man power to operate and requires less surface preparation. It can also load in various mining conditions,
has longer life, higher availability and can also do production by staying in the inclined terrain.
 However, it has low Manoeuvrability, no flexibility, is affected by climatic and watery
conditions, bank slides etc. Shovel can be used in strip mining method (loading, swinging and dumping of
overburden material into the adjoining excavated area by overcasting), in tandem operations (where 2 or
more shovels are used for rehandling of overburden successively), shovel pull-back operations (which is a
combination of a shovel and dragline for handling the overburden). Shovel is used for following major
operations:

1. digging a deeper cut

2. loading and overcasting and pull backing the spoil material
3. over burden removal in the contour mining in the hilly terrain
4. in over burden removal in open pit mining system
5. excavation in the face and loading on to trucks, bunkers or railway wagons
6. removal of top soil
7. construction of access roads and haul roads
8. opening up a mine by a box cut system etc.

# Selection of Shovels: The selection of a shovel depends upon the following major factors:

1. Requirement of daily production

2. Type and quantity of material to be excavated
3. Total or reserve, geotechnical parameters like
a. Thickness
b. Compressive strength
c. Tensile strength
d. Shear strength
e. Abrasiveness
f. Dip etc.
4. Bucket Fill Factor-larger shovel digs better than smaller one
5. Swell Factor
6. Working Cycle Time
7. Capacity of the haulage equipment
8. Availability of the electric power supply
9. Drainage Conditions.
10. Weight & Maximum lump size of the material to be excavated
11. Height of the bench and the height of the cut
12. Capital Cost of the Equipment
13. Operating and Maintenance costs
14. Availability of the spare parts
15. Facility of after sales services
16. Reliability of the machine
17. Climatic conditions & drainage system
 A shovel is identified by the capacity of its bucket. it is of two types

1. Rope Shovel: In a rope shovel, a bucket is rigidly connected to the end of the dipper stick, The dipper
stick can move forward or backward. It is supported by a pinion and is held by a cable to the bucket side.
The pinion and the rack on the underside of the dipper stick, is in constant mesh, which are attached in
the proper conditions, on a heavy structural boom. The lower end of the boom is pivoted with a
turntable, while the top end is held in the position, by means of a boom hoist cable. The superstructure
or the turntable carry the operators control room, mechanism for hoisting, crowding, swinging, and
travelling etc. of the shovel. The crawler mechanism supports the whole shovel unit and also provides
the travelling and steering motion of the shovel. Shovels have a bucket size ranging between 0.3m3 to
200m3. The main parts that are found in this machine, are as follows:
a. Bucket: The size of the Rope Shovel is determined by the capacity of its Bucket. Th greater is the
size of the bucket, the greater is the production.
b. Dipper Stick: The bucket is attached to the Dipper Stick, which is further attached to Boom. This
moves backwards and forwards, with the help of a Rack & Pinion arrangements. The bucket is
brought to the loading place, by operating the Dipper Stick.
c. Boom: The larger is the size of the boom of the shovel, the greater is the load, that can be bear
by it. The boom can rotate for an entire 360 degrees over the Chassis.
d. Machine House and “A” frame: This acts an Operator’s Cabin of the Machine, from where, the
entire machine is controlled. It has controls for the Hosting Drum and the Shovel Rope.
e. Crawler Chassis: The entire machine and its attachments, are attached to the Crawler Chassis,
and which can be entirely rotated 360 degrees, and can also be taken from one place to another.
The machine is travelled from the loading place, to the working place and vice versa, by the help
of it only.

The application of the machine, is as follows:

f. It’s used at places, where the loading of the material has to be done accurately.
g. It is also suitable at places, where the material has to be loaded onto dumpers, trucks or rail
wagons.
h. At places, where the loading material is harder than the material shaft. It is used in dry mines
only.
i. At places where the roof and floor conditions are strong.
j. At places, where the benches are sufficiently wide, for the installation of the transportation
arrangement.
k. It is also suitable for disposal of the overburden, out from the mine.

# Working: In order to load the bucket, the bucket is moved to the place, by the help of the dipper
stick & Hoist Rope. This process is called Crawling. The bucket is gradually raised upwards, by the
help of hoist rope, and it gets loaded. At this moment, the dipper stick is pulled backwards, and the
bucket is pulled out. This process is called Retracting.

Now, the whole machine, including the loaded bucket, boom and the
machine house, is brought over the dumper. This process is called “Swinging”. In order to load the
material onto the dumper, the operator operates a lever, through which, the bottom plate of the
bucket is opened, and the entire material falls down onto the dumper. The Bottom Plate gets
automatically closed. The machine is now swinged the entire way to the loading place, where the
process is repeated again. The lower is the angle of swing, the smaller is the bucket cycle time, and
the greater is the Production.
 Nowadays, the shovel with a bucket capacity smaller than 5m3 are not
manufactured. Generally, in India, the shovels with bucket capacities between 5, 10 & 20m3 are
used.

# Advantages: are as follows:

1. The rope shovel is able to load the material of the place, over which it stands.
2. As it is able to rotate 360 degrees, the Dumpers can be loaded at any place, in any condition.
3. It can be operated on both Diesel and Electric.
4. The lesser is the Angle of Swing, the lesser is the Loading Cycle Time, and greater is the
production.
5. It has a low requirement of Maintenance and its working is simple.

# Disadvantages: are as follows:

1. The teeth of this machine are deteriorated pretty quickly and have to be replaced frequently.

# Output of Shovel: are as follows:

S. Bucket Capacity & Dumper Size Production of Solid

No. Rock per Year
(In million cubic
meters)
1. 4-6 m3 With 35 tonne Dumper 0.87
With 50 tonne Dumper 0.90
2. 5 m3 With 35 tonne Dumper 0.95
With 50 tonne Dumper 0.98
3. 10 m3 With 85 tonne Dumper 1.98
With 120 tonne Dumper 2.08
4. 20 m3 With 170 tonne Dumper 4.09

2. Hydraulic Shovel: These shovels are slowly replacing the conventional rope shovels. Because of their
high productivity & higher efficiency, these are gaining much importance in the modern mining
operations. In these, the swinging mechanism is very similar to that of a rope shove, except the fact that
it is operated by Hydraulic Pushers, instead of Ropes and Dipper Stick. Moreover, the wrist action of the
bucket is possible in this shovel. For a same capacity of bucket, a hydraulic shovel is half in weight, as
compared to the rope shovel, due to which it is easier to move from one place to another, and the
Loading Cycle Time is also less.
These are also very reliable in their operations. The main operations in these
are done with the help of Hydraulic Pressure.
# Working Principle: The hydraulic pressure is generated by a prime mover, i.e., by a diesel engine
(most commonly used) or by an electric motor which drives number of, hydraulic motors, via a gearbox
used to perform most of the working functions. The hydraulic pump generates fluid pressure to the
valve bank which direct and controls the maintenance of fluid pressure in numerous hydraulic circuits
for the operation of various hydraulic units with higher efficiency. The hydraulic power operates all the
crowd mechanism, hoist mechanism, swinging mechanism, traction mechanism, etc. Presently hydraulic
excavator having bucket capacity wearing from 0.4 cubic meters to more than 34 cubic meters are
available in the market.
# Construction: It consists of the following main parts:
(i) Propelling Unit: This shovel is crawler mounted, with four geared wheels (2 in front and 2 in
back). The wheels are operated by electric or diesel operated motor, that further operates
the Crawler chains connected to them.
(ii) Operational Unit: The operational or functional unit of the shovel can be either electrically
or diesel operated. In case of electric, the unit is powered by 1100 K.W. power, through an
Electric Cable.
The Unit consists of two tanks, namely Hydraulic Tank and Accumulator
Tank. The former is used for Hydraulic Movements, and is filled with a mixture of 95% water
and 5% vegetable Oil. The latter is used for the creation of Pressure. The power is first
supplied to the motor, that causes it to start.
(iii) Control Unit: This unit consists of the Operator’s Cabin, that consists of almost all controls
and levers of the machine. This cabin is dust-proof and moisture-proof.
(iv) Swinging Unit: It is a special arrangement. With the help of this, the upper part of the shovel
(i.e., Operation Unit, Control Unit and Boom etc.) can rotate for 360 degrees, without
moving the lower portion (i.e., Propelling Unit). With this, the shovel can extract the OB or
Mineral on one side, and dump it on the other, without the requirement of moving entirely.
(v) Loading Unit: It consists of a boom, that is made of high-grade structural iron. This is moved
with the help of Hydraulic Power. The front portion of it consists of a bucket, that is used for
picking the broken ore or mineral. These buckets are of different capacities.

# Applicability: is as follows:

(i) It is used for small mining operations, such as small-scale mines, or patches
(ii) It is generally preferred for the removal of OB.
(iii) It is also preferred for Selective Mining.
(iv) It is also preferred for accurate loading of material.

# Advantages: are as follows:

(i) High productivity

(ii) Higher efficiency
(iii) higher machine availability
(iv) smoother and easier control
(v) possibility of application in both the pushing and pulling forces
(vi) smooth and controlled digging and sumping
(vii) faster loading cycle
(viii) being a crawler mounted machine, the grip during movement is greater.
(ix) ability of selective mining.
(x) less bulkiness and weight compared to a conventional shovel

# Disadvantages: are as follows:

(i) Maintenance Cost is more

(ii) Skilled and Efficient manpower are required for its operations
(iii) Power costs are more.
 (iv) It is difficult to keep the oil free from dust, inn dusty atmosphere of a mine.
(v) It cannot be used for very large-scale mining operations.
(vi) Movements are slow.

# Output of Hydraulic Shovel: is as follows:

S. Bucket Capacity & Dumper Size Production of Solid

No. Rock per Year
(In million cubic
meters)
1. 3-8 m3 With 35 tonne Dumper 0.95
With 50 tonne Dumper 0.97
2. 5 m3 With 35 tonne Dumper 1.09
With 50 tonne Dumper 1.11
3. 8-10 m3 With 50 tonne Dumper 1.95
With 85 tonne Dumper 2.13

# Various Attachments to Shovel: The shovels can be used with various attachment, the major ones
amongst which are as follows:

(i) Bucket: The capacity oof a shovel, is identified by the capacity of its bucket. The bucket size
ranging between 0.2 m3 to 200 m3 are available in the market. In India, the Buckets of 5, 10 and 20
m3 are commonly used. The greater is the size of the bucket, the greater will be the productivity of
the shovel.
(ii) Dipper Stick: The bucket is attached to the Dipper Stick, which is further attached to Boom. This
moves backwards and forwards, with the help of a Rack & Pinion arrangements. The bucket is
brought to the loading place, by operating the Dipper Stick.
(iii) Boom: The larger is the size of the boom of the shovel, the greater is the load, that can be bear by
it. The boom can rotate for an entire 360 degrees over the Chassis.
(iv) Machine House and “A” frame: This acts as an Operator’s Cabin of the Machine, from where, the
entire machine is controlled. It has controls for the Hosting Drum and the Shovel Rope.
(v) Crawler Chassis: The entire machine and its attachments, are attached to the Crawler Chassis, and
which can be entirely rotated 360 degrees, and can also be taken from one place to another. The
machine is travelled from the loading place, to the working place and vice versa, by the help of it
only.

# Comparison Between Rope & Hydraulic Shovel: is as follows:

S. Criterion Rope Shovel Hydraulic Shovel

No.
1 Initial Cost High Less
2 Machine Capacity High Comparatively Less
3 Flexibility Less More
4 Performance Great Comparatively Poor; cannot be
used in rough mining conditions
5 Working Space Requires more space Requires less space
6 Back Hoe Arrangement Cannot be installed Can be installed
7 Maintenance The cost is more, but is required Cost is less, but it is required more
less frequently frequently.
8 Bucket Movement The bucket can be only moved The bucket can also be moved in
in Up or Down Direction forward and reverse direction.
Moreover, the teeth installed in
 the bucket, can also be moved.
9 Working Cycle Time More Less
10 Reliability under Heavy More Less
Conditions
11 Steering Difficult Comparatively Easily

# Back Hoe Shovel: It is also known as a Pull Shovel or Drag Shovel. In this type of shovel, the bucket
face is towards the machine. It stands on the top of the bench, and excavates the ground below it. It is
very similar to the normal shovel. The boom is comparatively longer and the bucket is installed in the
reverse direction.

It just acts like a spade, but is mechanized and greater in size. Its efficiency is very
less as compared to a face shovel, and hence, it is used only in special conditions. It is comparatively
Expensive. Its boom is controlled by 2 hydraulic cylinders, and a bucket stick is connected with the
boom. The bucket is attached on the other end of the Bucket Stick, and it faces towards the machine.
The movement of the m=bucket is caused by the Bucket Cylinder. The capacity of the bucket ranges
between 0.38 m3 – 18 m3. It can excavate up to a depth of 4-8 m., in all the directions around it. Its
efficiency, is around 60-70% of that of a Dipper Shovel of the same size.

# Specifications: of the various models of it, manufactured by O&K, are as follows:

S. Criterions Models
No. RH30C RH40C RH75C RH120C
1. Backhoe Range (in m3) 1.8 – 4.7 2.0 – 5.6 3.6 - 10 6 – 18
2. Breakdown / Crowd 240 / 280 330 / 330 415 / 450 700 / 750
Forces (in kN)

# Application: of it are as follows:

1. It is suitable for the construction of Deep Long Drains, all around the mine, which are
known as “Garland Drains”.
2. It is used in opencast mines, where there is excessive mud and water.
3. It is used for removal of OB and for other shallow excavating cutting operations.
4. It is widely used for excavation of underground pillars, extracted by Opencast Method.
5. Mostly the Backhoes, ranging between 1.5 – 3.5 m3 are only used in the mines.

# Operating Parameters: are as follows:

1. Dumping (Discharge) Height (maximum): At a particular boom angle, it is the vertical distance
between the level (where the shovel rests) and the bottom most point of the bucket under the
tipping condition while the boom and all other accessories attached on boom are at their full
extended position under that condition (Fig 6.2(A)). Higher the boom angle higher would be the
dumping height for a particular shovel. A dumping height of around 46 m is common nowadays.

2. Cutting (digging) height (maximum)-At a particular boom angle it is the vertical distance
between the level (where a shovel rests) and the top most point of the bucket (tip of the teeth)
when the boom and all other accessories attached on boom are at their full extended position
under that condition (Fig 6.2(A)). Higher the boom angle higher would be the cutting height for a
particular shovel. A cutting height of 51m or more is not uncommon today.

3. Dumping (discharge) radius (maximum)- At a particular boom angle it is the horizontal distance
between the vertical swing axis of the main body of the shovel and the vertical centre line of the
 bucket when the boom and all other accessories attached on boom are at their full extended
position under that condition (Fig 6.2(A) Higher the boom angle lower would be the dumping
radius of a particular shovel. A dumping radius of 65m. or more is possible at present.

𝐴𝑐𝑡𝑢𝑎𝑙 𝑉𝑜𝑙𝑢𝑚𝑒 𝑜𝑓 𝑡ℎ𝑒 𝑚𝑎𝑡𝑒𝑟𝑖𝑎𝑙 𝑖𝑛𝑠𝑖𝑑𝑒 𝑡ℎ𝑒 𝑏𝑢𝑐𝑘𝑒𝑡

4. Bucket Fill Factor = × 100
𝑉𝑜𝑙𝑢𝑚𝑒 𝑜𝑓 𝑡ℎ𝑒 𝐵𝑢𝑐𝑘𝑒𝑡
It depends upon the degree of fragmentation of
mineral body, rocks or physical characteristics of the rock or ore, I.e., the size & shape of ore or
overburden rock, stickiness, watery condition, diggability, soft, loose & fragile characteristics and
also on the size of the bucket. The Bucket Fill Factor for the following materials, is as follows:
S. Material Bucket Fill Factor
No.
1 For easily digging rocks like sand, clay, loam etc. 0.75 to 1.0
2 Average Digging Rock like well blasted rocks 0.55 to 0.75
3 Difficult digging rocks like bulky, irregular shaped or rugged 0.35 to 0.55
rock, which is very difficult to scoop up

𝑤𝑒𝑖𝑔ℎ𝑡 𝑝𝑒𝑟 𝑢𝑛𝑖𝑡 𝑣𝑜𝑙𝑢𝑚𝑒 𝑜𝑓 𝑡ℎ𝑒 𝑠𝑜𝑙𝑖𝑑 𝑟𝑜𝑐𝑘 𝑖𝑛 𝑏𝑒𝑛𝑐ℎ𝑒𝑠

5. 𝑆𝑤𝑒𝑙𝑙 𝐹𝑎𝑐𝑡𝑜𝑟 =
𝑤𝑒𝑖𝑔ℎ𝑡 𝑝𝑒𝑟 𝑢𝑛𝑖𝑡 𝑣𝑜𝑙𝑢𝑚𝑒 𝑜𝑓 𝑙𝑜𝑜𝑠𝑒 𝑟𝑜𝑐𝑘 𝑚𝑎𝑠𝑠 𝑎𝑓𝑡𝑒𝑟 𝑏𝑙𝑙𝑎𝑠𝑡𝑖𝑛𝑔 𝑜𝑟 𝑙𝑜𝑜𝑠𝑒𝑛𝑛𝑖𝑛𝑛𝑔

It indicates the amount of increase in volume of a

rock mass in percentage after expanding due to loosening by digging action or by blasting. Poor
fragmentation due to blasting will increase the amount of swell since heavy amounts of voids
will be present in between the boulders.

6. Maximum digging radius- At a particular boom angle it is the maximum horizontal distance
between the tip of the bucket (tip of the teeth) and the vertical swing axis of the shovel (Fig
6.1(A)). Higher the boom angle, lower would be the digging radius. A digging radius of 10m or
more is quite common at present

7. 𝐵𝑢𝑐𝑘𝑒𝑡 𝐹𝑎𝑐𝑡𝑜𝑟 = 𝐹𝑖𝑙𝑙𝑎𝑏𝑖𝑙𝑖𝑡𝑦 × (𝑆𝑤𝑒𝑙𝑙 𝐹𝑎𝑐𝑡𝑜𝑟)−1

Fillability- Because of the slope line in front of the bucket
the struck capacity of the bucket is reduced to some extent depending upon the size of bucket
and the design. It is determined by the field observation and experience. The fill ability of the
bucket is expressed in percentage. It is always less than 1. Reciprocal of the Swell Factor is taken
into account for calculating the bucket factor.

8. Swing Factor- Standard cycle time of a shovel is based on its 90° swing for loading. This cycle
time will increase or decrease depending upon the increasing and decreasing angle of swing (a
shovel can generally rotate at 4 to 5 RPM). So, in such a case a certain factor is to be multiplied
to correct the shovel cycle time. This factor is known as the swing factor for which a standard
chart is provided by the manufacturer This swing factor is 1 at 90 degrees and it increases or
decreases depending upon the increasing or decreasing the swing angle respectively
corresponding to that standard 90-degree angle of swing.

9. Cycle Time- It is the total time taken by a shovel to complete one full cycle of operation starting
from the crowding operation into the face to swinging, dumping and again coming back to the
face for crowding operation.
 Total cycle time = Crowding (digging, loading, hoisting) time + swinging time towards the truck +
dumping time on the truck+ swinging back for crowding operation

In the shovel operation some amounts of delays are always present

which include the following:

a. Moving the equipment of proper positioning.

b. Delay due to non-availability of the haulage unit.
c. Floor cleaning by the bulldozer for the movement of the shovel
d. Dressing overhanging ledges in the high wall and also blending operation for quality
control
e. Improper spotting of the haulage unit and hence poor unloading.
f. Delay due to fatigue of the operator, unskilled operator, etc.
g. Due to poor blasting of rock when big boulders are formed.
h. Adjustments of the Boom Angle & other mechanical parts
i. Power cable shifting and damage of machine breakdown- both electrical or mechanical
units

Time is also lost due to fog, rain, tremendous hot waves,

cyclone, repair, maintenance & overhauling, shift changing, interruption of power etc.
𝐴𝑐𝑡𝑢𝑎𝑙 𝑊𝑜𝑟𝑘𝑖𝑛𝑔 𝐻𝑜𝑢𝑟𝑠
10. 𝐸𝑓𝑓𝑒𝑐𝑡𝑖𝑣𝑒 𝑂𝑣𝑒𝑟𝑎𝑙𝑙 𝑃𝑒𝑟𝑐𝑒𝑛𝑡𝑎𝑔𝑒 𝑜𝑓 𝑈𝑡𝑖𝑙𝑖𝑧𝑎𝑡𝑖𝑜𝑛 𝑜𝑓 𝑎 𝑆ℎ𝑜𝑣𝑒𝑙 = 𝑇𝑜𝑡𝑎𝑙 𝐻𝑜𝑢𝑟𝑠
× 100
𝑆𝑐ℎ𝑒𝑑𝑢𝑙𝑒 𝑆ℎ𝑖𝑓𝑡 𝐻𝑜𝑢𝑟𝑠−𝑀𝑎𝑖𝑛𝑡𝑒𝑛𝑎𝑛𝑐𝑒 𝐻𝑜𝑢𝑟𝑠−𝐵𝑟𝑒𝑎𝑘𝑑𝑜𝑤𝑛 𝐻𝑜𝑢𝑟𝑠
11. 𝐴𝑣𝑎𝑖𝑙𝑎𝑏𝑖𝑙𝑡𝑦 𝐹𝑎𝑐𝑡𝑜𝑟 =
𝑆𝑐ℎ𝑒𝑑𝑢𝑙𝑒 𝑆ℎ𝑖𝑓𝑡 𝐻𝑜𝑢𝑟𝑠
𝑆𝑐ℎ𝑒𝑑𝑢𝑙𝑒 𝑆ℎ𝑖𝑓𝑡 𝐻𝑜𝑢𝑟𝑠−𝑀𝑎𝑖𝑛𝑡𝑒𝑛𝑎𝑛𝑐𝑒 𝐻𝑜𝑢𝑟𝑠−𝐵𝑟𝑒𝑎𝑘𝑑𝑜𝑤𝑛 𝐻𝑜𝑢𝑟𝑠−𝐼𝑑𝑙𝑒 𝐻𝑜𝑢𝑟𝑠
12. 𝑈𝑡𝑖𝑙𝑖𝑧𝑎𝑡𝑖𝑜𝑛 𝐹𝑎𝑐𝑡𝑜𝑟 = 𝑆𝑐ℎ𝑒𝑑𝑢𝑙𝑒 𝑆ℎ𝑖𝑓𝑡 𝐻𝑜𝑢𝑟𝑠

o Dragline: A Dragline is an excavator which has a boom, one end of which is attached with the revolving unit
of the machine & the hanging end on the other side, carries a large sheave for a cable attached with a
bucket. The Dragline sits above the waste or Overburden, and excavates the material in front of itself, to
dump it on the spoil bank. It can dig through loose to mid-compact soil, as an intermittent excavator. In
India, the total population of the dragline is around 50, and the maximum capacity bucket size, is around 30
m3.
▪ System Of Working: The dragline may either be crawler mounted (most common), wagon mounted,
track mounted or walking type. The boom can move both vertically (from horizontal), 25 to 60
degrees (more common is 30° to 35°) and horizontally (0 to 187 degrees with the help of swing
mechanism) to perform the job. Rear end of the box shaped bucket whose one end is open, is
attached by means of two hoist chains with the help of pins. The other end of the hoist chains is
fitted with a dump sheave. To prevent the chains from rubbing on the side of the bucket a spreader
bar is attached with the chains before connecting with the sheave. The Drag Chains are attached in
the front side of the bucket, while the other end is connected with drag yoke. The Drag chains are
connected to the drag cable by dragline socket.

The boom is lowered and raised by the cable of the boom hoist. For
dragging the bucket towards the machine, one end of the drag cable is attached to the bucket & the
other end is connected to the bucket hoist, located at the foot of the boom. The bucket is filled up
by pulling or dragging the bucket against the loose material by the drag cable, and then it is hoisted
up by the hoist cable. Finally, it dumps the material directly over the spoil dump or the trucks or
railway wagons.
 ▪ Application: is as follows:
1. Gradient Flatter than 1 in 6
2. Seams shall be free from faults and other geological disturbances
3. Deposits with larger strike length (>2km.), so that frequent shifting of dragline from one side to
other side can be reduced.
4. The property should be large enough to ensure the life of about 25 years or more.
5. A hilly property is unsuitable.
6. Thick Seam (>25m.) are not suitable.
7. The dragline can be used for dredging.
8. It can be used for digging trenches, strip OB, clean & dig road-side ditches.
9. Also used for slope embankment, when it is handling mud.
▪ Advantages: are as follows:
1. Generally, the Draglines are used for direct handling and rehandling of overburden material during
overcasting, since it is the cheapest means of Overburden Removal.
2. It is also used to handle the soft and unconsolidated materials, blasted rocks or minerals, soil, top
soil, rehandling of ore or coal stockpile etc.
3. The Dragline can dig well below and above the level where it stands and has large flexibility in
operating conditions as compared to a shovel.
4. It can be used for stripping into the under-water, stripping of soft ore, sub-level operations,
removing a high cut and stacking in one operation.
5. It can also be used to work in muddy and unstable conditions, digging semi-consolidated soil, etc.
6. It cannot work on a steep grade on the uneven footwall condition.
7. Maintenance is cheap
8. It is not affected by Surface Bench Slide, Water seepage, etc.
9. Multi-seam Stripping is possible.
▪ Disadvantages: The Dragline has following demerits also-
1. If the blasted rocks are of large lump, the bucket is filled insufficiently and the bucket and drag rope
wear rapidly.
2. Used only for softer rock formation
3. Production cost is more as compared to the powered shovel,
4. Its efficiency is less as compared to a powered shovel,
5. Bucket fill factor is less as compared to the shovel and it requires well blasted/fragmented rocks to
be excavated
6. Lesser spotting ability
7. Lesser output than the powered shovel.
▪ Comparison With a Shovel: is as follows:
1. Bucket Fill Factor: Less than that of shovel by 12%.
2. Cycle time: more inn Drag Line by about 10%
3. Multi seam Project - shovel can do easily
4. Bearing pressure- DL-1 atm, shovel-3 to 5 atm
5. Floor conditions - DL require nearly horizontal
6. Reach- DL has longer reach.
7. Digging below level - DL can cut more below than a shovel.
8. Maintenance Cost- less in OL by 18%.
9. Frequency of failure- more in case of shovel.
10. Output- shovel's output is around 20% more than DL
11. Production cost- 10% less in case of shovel.
12. Efficiency- less in DL.
Failure frequency (dl*)
Bearing pressure (dl-1atm s-3_5atm)
Output (20 S)
Multi seam (s)
Efficiency (s)
Floor conditions (s)
Bucket fill factor (12 S)
Reach(dl*)
Cycle time (10 S)
Digging below level (dl*)
Production cost (10 S)
Maintenance Cost (18 DL*)
 Unit 2 (LO2)
o Bucket Wheel Excavator: The Bucket Wheel Excavator acts like a production Machine, and performs the task
of cutting the soft coal, and dumping it on the backside of the machine. It is an enormous machine, that is
one of its kind. It has a rotating cutting wheel in its from portion, for cutting the coal. The Wheel has a no. of
buckets, fitted with it. In India, it is being used only in Nevyeli Lignite Mine of Andhra Pradesh. This machine
is suitable for long range stripping of soft overburden rocks at a considerably lower cost although the
machine is costly having lower flexibility. The rate of production by the BWE varies between 100m3/h to
more than 11000m3/h. Machine Weight varies between 35t to more than 7000t & power around 200kW to
more than 7000kW.
▪ Application: The machine is nicely applicable in the following conditions:
1. Hard & Tough well fragmented blasted rocks with no or less boulders, having consistency of uniform
ground and bank condition.
2. Since it has a wide radius of excavation (around 40-90m.) with high & deep cut, the width of the
bench floor exposes more reserves and creates huge amount of space for the mobile equipment.
The slope of the pit is also very stable.
3. It can also be used for selective mining & thin seam mining.
4. For easy disposal of ore or overburden to a considerable distance above or below of its working
level.
5. It is a very efficient excavator for lignite, soft alluvium etc.
6. For reclamation of land.
7. It is used at places, which are free from any geological disturbances.
8. A large space is required for the operation of this machine
9. The deposit shall also be large enough.
10. If the capital cost of the mine is less, this machine can be used for providing Production at lower
costs.
▪ Construction: The main parts of a BWE are as follows:
1. Propelling Unit: The Bucket Wheel Excavator is a Crawler Mounted Machine, it is fitted with a large
number of chains, that are for the purpose of bearing the entire load of the machine.
2. Production Unit: As the Production Unit, The Bucket Wheel Excavator has buckets, that cut the soft
coal or material, and while the wheel rotates, the material from the buckets falls into the machine
boom. The bucket is designed in such a way, that its ends are conical for cutting into the coal. With
the movement of the boom, it can cut up to a height of 20m. It can also be used for the purpose of
Digging up to a depth of 3m. The size of the bucket and the diameter of wheel depends upon the
size of the machine.
3. Swing Unit: When the wheel rotates, the buckets cut into the soft coal and then discharge the
broken material in to a chute. The Bucket Wheel Boom can be moved upward or downward & in the
left & right direction. Similarly, the conveyor can also be moved. These movements are only known
as “Swinging”.
4. Transportation Unit: The BWE has two conveyors, which are as follows:
a. Bucket Wheel Conveyor: The coal from the bucket, first reaches into this conveyor and is
then transported to the discharge conveyor.
b. Discharge Conveyor: The Bucket Wheel Conveyor drops the broken material into this
conveyor, and through this, the material reaches to the final loading point, where it gets
loaded onto the trucks, rail wagons.
5. Control Unit: It includes the Operators cabin from where, all the operations of the machine are
controlled.

The main units of a Bucket Wheel Excavator,

are as follows:

1. Wheel Unit with Buckets:

2. Under Frame with Travel Mechanism
3. Superstructure with all machine housing along with motors, gearboxes & other mechanism, counter
weight structure, winches etc.
4. Swinging Mechanism
5. Digging Boom
6. Belt Conveyor
7. Swinging Cantilever with Discharge boom belt conveyor & chute
8. Luffing Arrangements
9. Suspension truss for digging boom etc.
10. Lubrication System.
▪ Operation: This machine is used for the extraction of soft materials. It is a crawler or rail mounted
machine. It has a long boom, with a Chain Conveyor Arrangement. A bucker wheel excavator (BWE)
has a wheel around 2.5m to 17 m diameter containing 6 to 18 number evenly spaced bucket
(capacity ranging from 0.04 m3 to 6.3 m3) around its periphery (Fig 6. 10). The buckets have teeth in
its front portion, which are made either of Manganese Ore or Tungsten Carbide (age-250 hours). The
series of buckets attached to the periphery of the wheel, dig into the mineral or softer rock mass and
cut the same when the wheel rotates from bottom to top (clockwise). The cut material is loaded by
the bucket and is discharged over the belt conveyor mounted on the same movable (both in

horizontal and vertical direction) boom via a hopper. The Length of the Belt Conveyor can be
changed. Moreover, it can also be swinged in Up or down direction, by the help of which, it can
discharge the material into any type of vehicle. Various types of buckets are available, which are as
follows:
1. Cell type buckets: has individual discharge chute to discharge material onto conveyor unit via
slope sheet or roll feeder or disk feeder.
2. The cell-less type bucket: discharges materials on to annular common guide chute permitting
free falling of material with higher wheel rpm and has higher capacity. The excavated material
from the bucket flows on to the conveyor unit via slope sheet or roll feeder or disk feeder.
 3. Semi-cell type buckets: are used for larger diameter wheel having advantages of highest bucket
filling and emptying characteristics.

The Boom can move in various directions, and according to

its length, the Caving Height of the machine is determined. The machine is supplied at 3300volts.

▪ Specifications: of the BWE being used in the Nevyeli Lignite Mines, are as follows:

S. No. Specification Details

1 Mine Neyveli Lignite Corporation Ltd.
2 Manufacturer B/W Bucket Wheel Excavator
3 I II
4 Wheel Diameter 8.4m. 10.5m.
5 Bucket Capacity 700 litres 1400 litres
6 Number of Buckets 9 10
7 Discharge per Minute 65 65
8 Wheel RPM 7.19 6.5
9 Cutting Speed 3.16m/s 3.57m/s
10 Theoretical Output 2730m3/h 5460m3/h
Wheel Drive
11 Main Drive 750kW 2*750kW
12 Auxiliary Drive 20kW 20kW
13 Maximum Cutting Height 20m. 27m.
14 Wheel Boom Length 26.3m. 31m.
15 Digging Depth 3.0m. -
16 Discharge Boom Outreach 35m. 35m.
Bucket Wheel Conveyor
17 Belt Width 1.8m. 2.0m.
18 Belt Speed 4.7m/s 4.5m/s
19 Conveyor Length 25.42m. -
20 Drive Capacity 200kW 2*200kW
Discharge Conveyor
21 Drive Capacity 200kW 200kW
22 Conveyor Length 37.6m. -
23 Travelling Gear 3 twin crawlers Arrangement of Crawlers
3-point support : 6ea
For fixed crawler group:
1*2ea
Steer Cable Crawler Group:
2*2ea
24 Travel Drive 6*45kW 600kW
25 Travel Speed 8m/min. 8m/min.
26 Service Weight 1309.93 t 2220t

▪ Advantages: are as follows:

1. It is the only machine in Surface Mining, that performs all tasks, including Production, Transporting &
Loading of materials.
2. It can easily cut through the soft coal
3. The bucket rotates continuously, which provides a Large Production.
4. Its Excavating Capacity is very great.
5. It can also be used for continuous operations.
▪ Disadvantages: as follows:
1. It is very expensive
2. It cannot be used for hard rocks.
3. It is difficult to operate
4. Its working is limited.
5. It requires a large space for its operations.
6. The timing taken for swing of the machine is great.
▪ Operations: There are two possible ways of Excavating a block, which are as follows;
1. Terrace Cut: The Terrace Cut is highly suitable for soft and easily diggable material because the
digging process is easiest to monitor under these conditions and is generally also easiest for the
excavator operator to manage. In this method, the machine cuts the bench, according to its height,
in a horizontal manner.
2. Drop Cut: In Drop Cut, the Bucket Wheel is lowered by the depth of one chip after each change of
chip, and if necessary, the machine is travelled back a little. This method is particularly suitable
harder materials to avoid oversized lumps.
In this method, the wheel is in constant contact with the ground, which means that the
superstructure cannot shake or swing. In addition to this, any oversize individual lump broken out of
the face are carried along the full height of the face before they are loaded. This means that there is
more time for each individual lump to be crushed against the face by the bucket wheel, as compared
to the Terrace Cut.

o Scraper: It is a diesel operated four-wheel drive rubber tyred tractor or a crawler tractor having a bowl
attached with a cutting blade at the bottom. The scraper can cut a thin slice of unconsolidated material, soft
rock or soil (thickness of the slice is approximately varying from 70mm to 250mm.) and can nicely be used
within a radius varying from 150m to 1500 m. However wet ground condition hampers the scraper
efficiency. It is very much flexible under wide varied conditions and can handle wide rate of production
ranging from very small to very high total tonnage. Scraper cut and load or simply load, haul, dump and
return back for recycling of operation
 Scraper can also be used for hauling blast fragmented hard
rocks or minerals. Scraper should be used for removal of top soil or unconsolidated rocks or minerals of
thickness 2m to 3 m. The speed of a tractor wheel mounted scraper is around 60km/h to 65 km/h. The
scraper may be used over a ripped or hard mineral or rock surface for loading and transporting purposes. It
is also used for levelling purpose. However, it requires a very good travelling way. It is unsuitable in hard
ground condition and sites having big boulders. The scrapers are basically divided into two categories:

1. Self-propelled type
2. Towed type.

Whilst operating on a bench, a scraper should not approach within 2 m. of

the bench crest. No scraper should be permitted to move backwards downhill when unloading. Scrapers
hauled by wheeled tractors should not be permitted to negotiate access roods having a gradient greater
than 15 degrees in the case of a loaded machine or 25 degrees in the case of an empty machine.

▪ Applicability: is as follows:
1. More efficient for flat gradient (4 degrees) but it can negotiate a gradient up to even more than 1.5
degrees.
2. Loading & Transporting earth over short distance (i.e., < 2miles)
3. Suitable for soft unconsolidated rocks
4. Size varies from very fine to more than 550mm. in good & dry ground conditions.
5.
▪ Operation:
1. Loading: For loading, the driver lowers the body of the scraper onto the soil & opens the gate. Next
a powerful bulldozer pushes against the back of the machine & pushes it in order to fill the body.
2. Transport: Once the body has been filled, the driver closes the gate & lifts up the body. The
bulldozer is no longer required as it only used during loading for more tractive effort.
3. Unloading: It takes place while the machine is moving. The driver opens the gate & activates the
ejector, which, with the help of a powerful jack, expels the materials form the front of the body
 o Ripper: Ripper is an excavating machine for loosening layers of soil, rock (usually medium thickness), using
the tractive force of a prime mover.
▪ Construction: It is a farmer's plough type steel shank (furrow) mounted with cutting tooth and
attached with a steel beam at an interval of 1m. to more than 2m. apart (average total numbers of
shanks are 4 but 6 teeth are also available) and the whole unit is attached at the rear side of the
crawler track mounted heavy-duty diesel operated tractor (in case of the tractor mounted type
ripper). The steel plough body attachment has a cutting tool at the bottom of it which dips into the
ground (to be ripped) at a depth varying from 0.4m. to more than 1.2m. depending upon the design
of the machine. It is done by the thrust applied to the hydraulic system. When the tractor starts
moving the friable soft to medium hard rock or mineral body breaks properly, which are loaded
thereafter with the help of a scraper or a loader or simply dozed with the help of bulldozer. The
spacing between the teeth should be optimum.
▪ Applicability: as follows:
1. This type of machine is suitable for ripping alluvial surface, soft rock, medium hard well stratified
rock, weathered sandstone or shale type rock, soft to medium hard lime stone, laterite deposits,
coal, well fractured hematite iron ore, etc.
2. It can also be used a to extract thin seams, where the thickness is less than 1.2m., where blasting is

Brittleness
Geological
Moisture
Fracture planes
Stratification
Consolidation not effective.
Weathering
Physio mechanical
Various strengths
The degree of rippability depends upon the brittleness of rock, degree of
stratification and lamination of rock, well defined fracture plane, moisture content in the rock, geological
disturbances like fault and other fractures, grain size, degree of consolidation and weathering, wet
condition, physio-mechanical properties of rocks like compressive strength, tensile strength, shear
strength, etc. Various types of rippers which are available in the market are as following-

1. Hinge type ripper- In this type the linkage hinge carries fixed position at the rear end of the ripper.
Various differential movements of the ripper teeth in the vertical direction due to the up and down
movement of the ripper cause the change of the teeth angle to meet the varying rock conditions.
 2. Parallelogram type ripper- In this type of ripper the linkage carrying the beam and the shank has the
constant teeth angle regardless of the teeth depth and have excellent penetration characteristics.
The ground clearance between the tracks and the shank of the ripper and the raising height of the
ripper is more which enable to rip more blocky lumps and ease out the checking of shank tip
respectively.
3. Impact ripper- In this type of ripper the engine power is converted into hydraulic power to impact
rapidly to a specially made ripper shank. The energy obtained from those impact pulses transmitted
into the ripper tip causing heavy fracturing and increasing the penetration depth into the rock or
mineral. It also reduces the draw bar pull. With this type of machine, ripping of hardest rippable rock
is possible.

o Bull Dozers: It is a diesel operated tractor mounted machine having a dozer (pusher) blade attached to the
front of it. The machine is either wheel or crawler track mounted having hydraulic arrangements to raise the
blade up and down. The concave straight blade's mould board is attached with cutting edge to dig into the
soft rock or soil surface and doze the loose cut rock or mineral body. These blades are generally light to
heavy in weight, depending upon the condition of the rocks. In conventional type of bulldozers, the blade is
kept in right angles to longitudinal axis of the tractor. The angle of blade is generally fixed, but in some cases,
it maybe varied. A dozer can dig more than 1.5m. below the ground. The travelling speed of it ranges
between 12-13 kmph.
1. The machine is also used for land preparation, cleaning, construction and maintenance of haulage roads,
benches etc.
2. It also serves as auxiliary back up service machine for piling up the blasted rock mass for shovel and
dragline.
3. Bulldozing the waste pile, making proper slope in the bench, cleaning up of bushes, grabbing, salvaging
and shifting of machines and materials like switches, cables, pipes, pumps, slid mounted auxiliary
machine units, etc.
4. They are also used for pulling heavy vehicles into the workshop for proper repair and maintenance when
they are under breakdown for pushing coal into ground bunker.
5. Used for reclamation of land, levelling, spreading and compacting of material in the dumping yard etc.

The bulldozer is most suitable equipment to handle the rock mass under the following conditions-
 1. To doze fine to large range blocky rocks or minerals
2. To work within the operating radius of 150 m
3. In all ground conditions (rough terrain, soft or wet ground)
4. It can also work at a gradient of 13 to 14 degrees. It is very much flexible under various ground condition
and has a high degree of mobility.
5. If the machine is used for hauling purpose its production rate is very low.
6. Dozers are also used for relocation of stockpile, spreading waste material on dumps, dozing of
supplementary material, pioneering and maintenance of roadways, etc.

During selection of a dozer, the following criteria shall be

considered:

1. Hourly production requirement

2. Number and size of equipment required
3. Attachment required (e.g., ripper) for special purposes
4. Crawler track or rubber tyre mounted machines
5. Suitability of machine in various topographic conditions
6. Site specific conditions and environment
7. Operating method
8. Requirement standard of maintenance
9. Requirement of skill of the operator
10. Performance statistics of the equipment in the other mines
11. After sales service
12. Availability of indigenous equipment and spare parts.

o Surface Miners: The Continuous Surface Miners are basically a cutting machine used in surface mines. It cuts
consolidated soil & semi-solid rocks without drilling & lasting. The cut material is pre-crushed, which makes
it suitable for belt conveying, loading & transporting.

These machines have been developed during the 1970s, primarily

for the extraction of thin seams (not less than 2.5cm thick) and selective removal of thin dirt bands present
in the seam. It was developed as a solution to fly rocks & cost-effective and quality-controlled mining. It is

basically a blast free method of mining.

▪ Applicability: is as follows:
1. It is widely used in projects where drilling & blasting is prohibited.
2. It can be conveniently used for mining of coal & Overburden.
3. This machine can cut the slices up to 10cm.
4. It is suitable for coal seams, not having geological disturbances like faults & folds & shall have a
minimum area of 200m*60m. for convenient Operation of the machine.
5. It can be used for a gradient up to 1 in 4 (best suited 1 in 9)
6. It is best suited for rocks having compressive strength of around 60 MPa.
7. It is suitable for less abrasive rocks
8. Non-Sticky materials
9. Thin seams with thin dirt bands
10. Sized material productions
11. Used for selective mining.
12. Mining of Harder Minerals
13. Mining of thin seam deposits
14. Creating channels and drainage ditches.
15. Moving conveyor units with an attached side boom
16. Road construction and maintenance
17. Mining of residual minerals
18. Digging Exploratory trenches.

▪ Types: The following are the major types of Surface Miners available in the market:
1. Bucket Wheel Type: Cutter picks are fitted on various bucket on the cutting drum, which is situated
inn the front of the machine. The machine can exert huge force in cutting both hard coal and rock.
This type of surface miner is basically used for mass production.
2. Milling Type: Cutter picks are fitted on drums and depth of cut is up to 600 mm (max). Production
capacity is low and can be used mainly for selective mining.
3. Boom Type: Cutting drum fitted with picks is mounted on a boom, like road headers. Due to longer
boom the cutting force of the machine is limited.
▪ Benefits: are as follows:
1. Mining is done without the blasting, hence there is no need of drilling. Moreover, there is:
a. No Vibrations
b. No Fly Rocks
c. Reduced noise & Dust development
2. In-situ Crushing is feasible & further sizing before conveying on to the truck is possible.
3. They have a unique conveying system, by which, the mined material is directly loaded on to the
dumper/truck.
4. The Haul Road preparation and maintenance in deposit are reduced. Sometimes, it is even not
required.
5. Production of small particle sizes directly during the mining process.
6. Selective Mining can be done with this machine effectively.
7. It replaces equipment for drilling, blasting, loading & auxiliary works, i.e., secondary blasting.
8. Low investment cost in comparison to equipment used in conventional mining.
9. Low operating costs due to less equipment & less personnel.
▪ Limitations: Are as follows:
1. This machine has limited application in seams steeper than 1 in 4
2. This machine cannot serve the purpose of selective mining in seams having dirt bands less than 5cm.
thickness only.
3. The performance of this machine is also reduced with the presence of geological disturbances in coal
seams, i.e., presence of faults, particularly in strike direction which limits the dimensions of the
working bench for surface miners.
▪ Construction: The Surface Miner mainly consists of 5 units, which are as follows:
1. Mainframe: It consists of the operator’s cabin, supporting frames, boom counter weight, slewing
ring, etc.
2. Conveying Unit: It consists of two stage conveyors, one carries the materials from the cutting drum
& discharges on the boom conveyor, which in turn conveys it onto the trucks.
3. Drive Unit: It consists of a diesel operated engine, whose power is transmitted to the cutting drum
via robust belt drive. And it consists of hydraulic system to operate other units.
4. Crawler Unit: Based on the design, it may be single or double pair crawler.
5. Cutting Unit: It is the most important component of the machine the selection of the right cutting
tool is essential for good performance & high life, since, it constitutes the major portion of the
operating cost of the machine. The cutting depth can be regulated by automatic levelling system
mounted on the machine, the pre-selected cutting depth is either maintained automatically or
manually.
▪ Classification: Based on the cutting principles & position of the drum, the surface miner is classified
in the following different types:
1. Machine with Middle Drum Configuration: In these types, the cutting drum is positioned below the
machine between the front and the rear crawlers, cutting picks are fitted on a drum and having the
depth of cut 500-600m.
2. Machine with Front Boom Cutting Drum: In these types, the cutter drum fitted with picks, is
mounted on a boom, like road headers. Due to longer booms, the cutting force of the machine is
limited.
3. Machine with Front Cutting Wheel: In this type, several bucket wheels are arranged in tandem and
cutter picks are fitted on lip of each bucket to form the cutting drum, which is positioned in the front
of the machine. It can exert huge force in cutting both coal and rock as the bucket wheel drum is
close to the machine.
▪ Working Principle: The surface miner is crawler-mounted machine having a cutting drum located
b/w two sets of crawlers and position -ed at the centre of the machine. The drum is lowered and
 raised by hydraulic system with powerful hydraulic motors thereby varying the depth of cut. The
material cut is loaded onto the primary and secondary discharge conveyors for loading the same on
to the loading/transporting equipment. The rear crawler travel at lower level then the front crawlers
to adjust to the required depth.

Dust Suppression is ensured by means of water

spray arranged on the cutting drum, which also serves the dual purpose of cooling of picks, thus
prolonging their useful life. The operator in cabin controls the speed, position of the belt conveyors
for the proper loading into the tippers. The cutting drum is followed by the scraper blade, which
gathers any material left on the floor.

o Motor Grader: A grader, also referred to as a road grader or a motor grader, is a construction machine with
a long blade used to create a flat surface. A motor grader should never be used for a site having less than
300m. in length & width less than 12m., since the turning of a grader, is a very tedious task.

▪ Construction: It is pneumatic wheel driven equipment; the back portion is the main body mounted
over four larger pneumatic tyre wheels & all control & driving arrangements are incorporated in it.
The front portion is mounted by two smaller pneumatic tyre wheels & connected with the rear main
body by a cross braced frame. A circle having geared teeth is suspended from cross braced frame. A
grading blade is attached with the circle, which can rotate to change the blade angle, depending
upon the condition.

A grader moves forward, while in operation, but it can also be

reversed or the direction of the blade may be reversed for doing spreading & smoothing the job. The
blade can be tilted in any side or lifted up or can be tilted vertically to dig the floor. It can also be
shifted to any side to a considerable length.

▪ Uses: are as follows:

1. For grading & levelling surface land
2. Light Cutting
3. Planning of bench berm & haulage road
4. Cleaning of spillage rocks or ore boulders over the haulage roads etc.
o In Pit Crushing System: In response to the increasing truck haulage costs and in the face of increasing
tonnage to be moved, two large U.S. copper mines responded in late 1960s and early 1970s by installing
high-capacity belt conveyor systems with large, permanent type, gyratory in-pit crushers. The systems did
not eliminate the trucks all together, but restricted their use only within the pit from the working faces to
the crusher, and the crushed material thereafter was transported to the plant (ore) and waste dumps
(waste) through belt conveyors.

Duval Corporation began installation of the first movable in-pit crushing and conveying
systems in 1981 at its Sierrita Copper and Molybdenum Mine near Tucson, Arizona, and the operation of the
system began in 1982 December. Operation of another similar system also began in 1982 at Sishen's North
Iron Ore Mine in South Africa. Since then, in-pit crushing and conveying systems has been adopted in many
surface mines, both coal and non-coal, throughout the world. In India, in-pit crushing and conveying systems
have been installed in surface coal mines at Piparwar Opencast Project (CCL), Ramagundam-II Opencast
Mine (SCCL) and Padmapur Opencast Project (WCL), and in Limestone mine at Jhinkpani (ACC).

▪ General Applicability of In-pit Crushing and Conveying Systems: In-pit crushing and conveying
systems may be used to transport ore (i.e., pay mineral) to the processing plant and/or overburden
(waste) to the waste dumps from the pit. In most of the cases, the pay mineral (which normally
requires primary crushing in its preparation process) can be transported by belt conveyors at lower
cost than by trucks. But for overburden conveying, the necessary preparation and dumping costs are
a considerable extra that requires to be offset by savings in trucking and truck haulage costs. In
general, only when large quantities of overburden are to be transported over fairly long distances at
considerable lifting heights, application of crusher- conveyor-stacker system may be justified.
▪ Types of In-Pit Crushing System: The various in-pit crushing systems may be classified broadly into
three categories Stationary/Permanent in-pit crushing system, Relocatable/Movable in-pit crushing
system and Mobile in-pit crushing system-
1. Stationary/Permanent In-pit Crushing System: In this type of systems, large capacity crushers
(separate crusher for ore and for waste) with proper foundation are installed in permanent
locations at or near the pit floor level. The material, excavated and/or loaded by shovel, from all
the working faces (of same type, i.e., ore or waste) are transported to the respective crusher by
trucks, and thereafter the crushed material is transported to respective location (ore plant or
waste dump) by stationary conveyors.

The crushers are usually positioned so as not to create any

inconvenience in the future extension of mine working and to ensure that uphill movement of
loaded trucks carrying material from the working faces to the crusher is minimized. However,
with the advancement of mine workings, the truck transport distance increases with time as the
faces moves away from the crusher location. This system is most suitable in case of surface
mines mining flat ore bodies so that –
 1. In course of operations the pit floor depth remains essentially same,
2. Large quantities of ore to be transported over a long distance that may or may not be
associated with considerable vertical lift.
3. The benches on one side of the pit (or at least a part of it) have reached their ultimate
position and became stationary.

2. Relocatable/Movable In-pit Crushing System: In such systems, the crushers used are of medium
capacity, and they that do not require extensive foundation work as required by stationary in-pit
crushers. In a mine, depending on the mine specifics (the number of working faces and their
relative location, material types to be handled total tonnage of ore and of waste to be handled),
there can be one or more such crushers installed. Generally, each crusher serves a group of
working faces excavating same type of material (ore or waste), and the crusher is placed at a
location such that the overall truck haulage distance from the working faces it serves is
optimized. The crusher, following the advance of working faces it serves, is relocated/moved to
the new load centre at appropriate time intervals.
In this system, also, the material excavated and/or loaded
by shovel is transported by trucks from the face to the crusher. The haulage distance is kept as
short as practicable and, if possible, uphill haulage is either avoided or restricted to a minimum.
The crusher discharges onto relocatable conveyor (that is relocated or lengthened in conjunction
with the relocation of the crusher) which in turn transfers the material to the stationary
conveyor. This system is suitable for a mine where –
1. The production is medium to large
2. The working faces are distributed in number of benches
3. A stationary in-pit crushing system is not suitable.

A relocatable in-pit crushing system has been installed at Ramagundam-II

Opencast Mine of SCCL. Altogether 3 numbers of double roll type crushers for overburden and
one double roll type crusher for coal have been installed. After crushing, the coal is conveyed to
CHP via coal conveyors and overburden is transported to overburden dumps via overburden
conveyor. In the dump area, the material is transferred to shiftable dump face conveyor for
subsequent dumping via tripper car and spreader. There are 3 numbers of tripper car and 3
numbers of spreader for dumping the overburden suitably at the dumpsites.
 3. Mobile In-pit Crushing Systems: In such a system, the crusher is fully mobile and is located at
vary near the face. The crusher is fed directly by an excavator (Rope shovel or hydraulic shovel or
front-end-loader). The crusher discharges either directly or via a belt wagon (or conveyor bridge

in some cases) onto the shiftable face conveyor. Finally, the material is conveyed to the --
respective destination (ore plant for one and waste dumps for waste) through a network of belt
conveyors. The crushers are generally of small capacity sufficient to handle the production from
the face where it is located. This system is applicable for mines where only a few fast-moving
working faces exist and the excavators stay permanently at the same face.

A mobile in-pit crushing system is in operation at Piparwar Opencast Project of CCL.

The in-pit crushing is done only for coal excavated at the working face of lowermost coal seam
(Lower Dakra), and the crushed coal is transported directly to the coal washery located near by
the mine via a system of belt conveyors. The components for the system are 25m³ rope shovel,
2800 TPH mobile roll crusher unit, belt wagon, shiftable face belt conveyor and other conveyors
(shiftable and stationary).

▪ Advantages: are as follows:

1. Improved safety due to decreased mobile vehicle usage
2. Reduced greenhouse gas emissions & Noise levels & Dust emission.
3. Reduced manpower requirements of between 40 – 60% due to higher automation
4. Reduced spare part and maintenance requirements
5. Reduced bad weather downtime
6. Mining operation will be continuous.
7. Improvement in the equipment availability and utilization
8. Elimination of fleet of dumpers will reduce diesel consumption along with saving in lubricants.
9. With adoption of the technology, it will be possible to work deep deposits.
10. High system availability.
11. Less operational expenditures (OPEX).
12. IPCC lends itself to easy automation.
13. Lower maintenance cost.
14. Highly reduced road preparation.

▪ DISADVANTAGES: are as follows:

1. The initial cost of system is normally higher than that of the truck haulage system, because the
complete conveyor and crusher are bought to start production whereas the truck fleet can be
bought in stages to set up production.
 2. The mining operation is completely dependent on availability of the conveyors. This availability
is over 95% but a shutdown of one belt can stop the entire production.
3. Relocation of the crusher and extension of the conveyor is expensive and requires a shutdown of
the mining operation for a period from 2-3 days.
4. Material must be crushed to a size of minus 250 mm before loading onto the conveyor.
5. IPCC and shovel do not operate together.
6. In-pit crushing required for conveying (hard rock) even if not needed (overburden).
7. Less flexible in mining layout & in capacity.
▪ Comparison: is as follows:
 Unit 3 (LO1)
Explosives
o Definition: An explosive is either a solid, liquid or mixture of compounds which when subjected to a
combustion (flame), detonation (sudden shock) or heat, undergo themselves to a rapid instantaneous
change into a large volume of gases at high pressure & temperature. The combustion or detonation process
initiate the chemical reaction among the compounds in the explosives. The released chemical energy, is
converted to heat and mechanical energy during blasting and the released high pressure breaks the rock.
The explosives should be very safe and sufficiently insensitive for any condition of handling & storage,
possess adequate strength, VDO & density, good water resistance property, fume characteristics and storage
properties, etc. Combustion is the process of a very high rate of burning. Detonation is the process of
propagation of a violent shock wave through a cylindrical charge initiating an instantaneous & high degree of
chemical reaction & producing sufficient quantity of energy for the propagation of shock waves. If
detonation is not there, then there will be deflagration in the burning. A deflagration moves very slowly to
produce shock waves sufficient to fracture rocks. Under sufficient detonation rock shatters whereas in weak
deflagration process, rock displaces. An explosive contains the following ingredients, beside the explosive
substance:
1. Oxidising Agents: Sodium Nitrate, Ammonium Nitrate etc.
2. Combustible Substances: Charcoal, Woodmeal, Sulphur etc.
3. Stabilizers: Calcium Carbonate, Magnesium Carbonate etc.
4. Sensitisers: Metallic Powder (Aluminium Powder), etc.
5. Flame Depressant: Sodium Chloride etc.

o Properties: The following properties of an explosive are of importance:

1. Strength: This is a measure of the amount of energy released by an explosive during blasting and hence
its ability to do useful work. The relative strength or power of an explosive is given by the term weight
strength in the case of explosives manufactured by ICI India Limited and a few other explosive
manufacturers. The weight strength, in the case of ICI explosives, indicates the strength of any weight of
explosive compared with the same weight of Blasting Gelatine which is taken as standard because it is
the most powerful commercial explosive manufactured by ICI. The weight strength of Blasting Gelatine is
100. At present, Blasting Gelatine is not in the regular manufacturing range of ICI India Limited. The
Ballistic mortar is calibrated initially with standard Blasting Gelatine and subsequently the weight
strengths of other explosives are determined with respect to the above calibration. Indian Oxygen Ltd.,
and Indian Detonators Ltd. do not use the term weight strength to indicate the relative strength of their
explosives.
The strength of the explosives are expressed as:
a. Absolute Weight Strength: It is the measure of absolute amount of energy available in unit mass
of explosives. It is expressed in kcal/kg. It is determined Theoretically.
b. Absolute Bulk Strength: It is the measure of absolute amount of energy available in unit volume
of explosive. It is expressed in kcal/kg. It is determined theoretically. It is the product of AWS and
density of the explosive.
c. Relative Weight Strength: It is a measure of energy available in a given weight of explosive
compared to an equal weight of ANFO.
d. Relative Bulk Strength: It is a measure of energy available in a given volume of explosive
compared to an equal volume of ANFO.

2. Velocity of Detonation: It is the rate at which the detonation wave passes through a column of explosive
and this is of considerable importance as the shock energy of detonation increases rapidly with this
velocity. It is expressed in m/s. Most of the high explosives, permitted explosives and slurry explosives
 used in the mines have a velocity of detonation ranging between 2500 and 5000 metres per second. For
high explosives which are used as boosters, the V.O.D. is high, e.g., O.C.G. 6000 m/s; Primer 7000 m/s.

It depends upon a number of factors including Explosive

formulation, charge diameter, degree of confinement, explosive density, etc. Ideal explosives are those
explosives, the VOD of which does not depends on their diameter, e.g., PETN. Non-ideal explosives are
those explosives, the VOD of which depends upon their diameter, e.g., ANFO. Most of the commercial
explosives shows non-ideal behaviour. It should be noted that the basic principle of detonation is more
intimate the contact between the oxidizer and fuel, the higher is the V.O.D. It should be noted that the
basic principle of detonation is more intimate the contact between the oxidizer and fuel, the higher is
the V.O.D.

3. Density: The mass per unit volume of an explosive is called its density. The density is important when
selecting an explosive for a particular use. It is also important property because explosives are
purchased, transported, stored and used in weight basis. It is expressed as g/cc or kg/m3. The density of
the commonly used commercial explosives varies from 0.8 to 1.5 g/cc. Density of an explosives depends
upon the density of the ingredients used to make it. It decides the amounts of explosives that can be
poured into a blast hole.
 With a high-density explosive, the energy of the shot is concentrated-a desirable
feature in tunnelling and mining operations in hard ground. On the other hand, when the output of lump
coal from mine is important, it is advisable to use a low-density explosive, which distributes the energy
along the shot hole. When an explosive is to be used in watery blast holes, its density should be more
than the density of the water present in the blast hole, otherwise it will float on the water.
4. Loading Density: The weight of explosives per meter of blast hole.

5. Water Resistance: Explosives differ widely in resistance to water and moisture penetration. Some
explosives deteriorate rapidly under wet conditions, but others are designed to stand water long enough
to enable the work to be done. When blasting is to be performed under wet conditions, a gelatinous or
slurry explosive3 should be used. The higher the nitro-glycerine content of an explosive, the better is its
water resistance qualities.

6. Sensitivity: An explosive is required to be insensitive to normal handling, shock and friction, but must
remain sufficiently sensitive to be satisfactorily detonated, and capable of propagating satisfactorily,
cartridge to cartridge and even over short gaps such as may occur in practice. It is expressed in the
following terms:

a. Initiation Sensitivity: of an explosive product is defined by the input energy necessary to cause
the product to detonate reliably. It is also called minimum booster rating or minimum primer
requirement.
b. Propagation Sensitivity: is a measure of the ability of an explosive to support detonation that
has already been created.
c. Gap Sensitivity: is the length of the air gap between cartridges in a single row.
d. Sympathetic Detonation: is defined as initiation of an explosive charge without a primming
device by detonation of an adjacent charge.

7. Fume characteristic: Explosives which are to be used where ventilation is restricted must produce a
minimum of harmful gases in the products of detonation. Slurry explosives and AN based explosives are
preferable to the NG based ones.

8. Detonation Pressure: It is the near instantaneous pressure derived from the shock wave moving through
the explosive column. When initiating one explosive with another, the shock pressure from the primary
explosive is used to cause initiation in the secondary explosive.

9. Cohesiveness: It is the ability of the explosive to maintain its original shape.

o Classification: as follows:
1. On the basis of speed: Explosives are grouped in two types depending upon the speed with which
explosive effects is produced:
a. Low explosives: Gunpowder is a common example of a low explosive. When a low explosive is
blasted the process of oxidation of the constituent substances is propagated by rapid
combustion from particle through the mass of the explosive and the of explosion is relatively
low. A low explosive is fired by ignition or a flame. The effect of Explosion energy and velocity of
propagation of flame is low in this case. Both the pressure & the temperature of the released
gases are built up gradually.
b. High explosives: High explosive always contains an ingredient which is explosive in itself, at least
when sensitised by proper means. A high explosive explodes when a violent shock is applied to it
with the help of a detonator; the process of oxidation does not proceed from particle to particle,
but is instantaneous and the constituents react with high velocity. High explosives therefore
produce a shattering effect. The effect of explosion energy & the velocity of detonation of this
type is very high. Both the pressure & the temperature of the released gases are built up
instantaneously. So, within its own volumes, the explosive compound converts into gases at a
very high temperature & pressure. Ammonium nitrate, nitro-glycerine, T.N.T., special gelatine,
slurry explosives, etc. are high explosive.

The terms "low explosives" and "high explosives" should not

be confused with the terms "low density explosives" and "high density explosives".

2. According to the Explosives Rule: Under the Explosives Rules, explosives & their accessories have been
classified under the following classes:
a. Class I – Gun Powders
b. Class II – Nitrate Mixtures (like ANFO, Aqua dyne, Ener gel, GN-I, Go dyne, Perm dyne, Perm
flow, Power Flo, Powe rite, Super dyne, Super gel, Toe blast etc.)
c. Class III – Nitro Compounds
i. Div. 1: Blasting Gelatine, Special Gelatine, OCG, Ajax-G, Viking-G, Soligex, Unisex-G etc.
ii. Div. 2: Gun Cotton, PETN, TNT etc.
d. Class IV – Chlorate Mixtures
e. Class V – Fulminate
f. Class VI – Ammunition
i. Div. 1: Safety Fuse, Igniter Cord, Connectors, Electric Lighters, etc.
ii. Div. 2: Cordtex, Detonating Fuse, Plastic Igniter Cord, Fuse Igniters etc.
iii. Div. 3: Detonators, Delay Detonators, Relays etc.
g. Class VII – Fire Works
h. Class VIII – Liquid Oxygen Explosives (LOX)
o Booster: For effective detonation of some slurry explosives and AN-FO mixture such as GN-1, use of a high
detonation-velocity booster is necessary. ICT India Ltd. manufactures a booster with the trade name
"Primex" which is a mixture of PETN and TNT. It is water resistant, has a velocity of detonation of 7,000
m/sec, weight strength as 82 and it can be detonated by a detonating fuse or, a detonator. The booster
manufactured by IDL Chemicals Ltd. is marketed by the trade name "Pentolite" which has a sp. gr. of 1.55 to
1.61. Compared with normal explosives boosters are quite costly. Primex is cast in cylindrical pellets
provided with two longitudinal holes for threading on to a down line of detonating fuse. For priming, a
detonating fuse is threaded through the two holes in the Primex pellet and a knot tied at the top. This
assembly is then inserted into a cartridge of slurry and its mouth re-tied by a wire. After lowering the primer
cartridge, are freely dropped down the hole.

A cast booster is not a substitute for the explosive charge; it may be compared with
a very powerful detonator of large size and is preferred for deep large dia. blast-holes in opencast mines.
During use a cast booster is knotted to the detonating fuse for placement at the bottom of the blast-hole
and additional boosters are threaded in the same detonating fuse so that their positions coincide with the
level of hard rock when the fuse is in position in the blast-hole. After lowering the booster by the detonating
fuse, AN-based site- nixed or plant-mixed slurry is poured or pumped in the hole. As the slurry is pumped in,
special ingredients are added to it at intervals to cover the hard-rock portion

Gelatinous NG-based explosives having high NG content like OCG or special gelatine, can be
used as a booster or primer explosive.

▪ Pentolite Boosters: These are manufactured by IDL Chemicals Ltd. and do not contain NG or other
headache-causing ingredients. These make efficient primers and boosters that detonate at a very
high velocity and temperature. These qualities, together with high density produce greate
detonation pressure. These can be stored for long periods of time, as they contain no liquid
ingredients and can be operated easily without the use of prickers. They are equipped with 2 holes
for easy initiation with detonating cord or detonators. A PETN based booster in a shot-hole occupies
only 1 to 2% of the total explosive charge but a high explosive like OCG, if used as a booster, occupies
15-20% of the total explosive charge.
no ng
High velocity and temperature this high pressure
Longer life ✓ Applications: Pentolite Boosters can be used for effectively initiating cap-
No liquid ingredients hence easy use
Detonating cord or detonator insensitive chares in a borehole, at any predetermined point in a coloumn of
explosive charge.
✓ Specifications: are as follows:
▪ Density 1.55 to 1.61 gm/cc
▪ Velocity 7800 metres per second
▪ Shock Energy 237 calories/gram
▪ Bubble Energy 470 calories/gram
▪ Detonation Pressure 240 kilobars
▪ Examples: Toe blast, KELVEX-800 etc.
▪ Example of Cast Booster: are as follows:
1. Primex booster: manufactured by IEL: It is used for non-cap sensitive explosive and is water
resistant. It is initiated by the cordtex detonating fuse and detonate at about 7000 m/s. Sizes
available are of the range of 23, 37, and 50 mm dia, with weights of 20, 100 and 250 gms.
respectively.
2. Avimex cast booster: manufactured by IGEL has a high VOD, high density and high strength and
contains a mixture of PETN and TNT for initiating and bosting non-cap sensitive explosives such
as NCN-400, NCN-500, NCN-600 including ANFO and Emulsion Explosives. It is available in 100
gms, 250 gms and 500 gms sizes. Density: 1.6+0.5 gms/cc, VOD: 7000 + 500 m/s.

o AN-Fuel Oil Explosives (ANFO): Ammonium nitrate, mixed with diesel oil, is used on a large scale for
blasting, in the quarries of coal and metal mines. The most effective and oxygen-balanced explosive mixture
is one with 5 to 6% of diesel oil (by wt.) The 6% Fuel Oil is important to raise the heat of explosion from
1.5KJ/g for AN Prills alone to 3.8KJ/g for ANFO. It has a sp. gr. of 0.8 to 1.0, wt. strength of 75-80 and velocity
of detonation at 3500 m/sec. In the dry season, 7 litres diesel for 100 kg of AN suffice but in wet season, the
quantity should be increased to 9 litres. Diesel oil in excess of 8% tends to lower the sensitivity of ANFO to
initiation.

The mixture causes irritation of the skin and the workers should, therefore,
wear gum boots and rubber hand gloves. The mixing should be done with wooden shovels avoiding contact
with iron. The mixture is safe to handle and without formation into cartridges can be mechanically loaded
into blast-holes. Where ANFO consumption is heavy, stationary ANFO mixer similar to the concrete mixer,
may be placed at a centrally selected site. In case of a Pneumatic ANFO loader, an electric detonator should
not be used unless steps are taken to prevent premature initiation due to static electricity

As the mixture cannot be initiated in the normal way by a detonator, it is necessary

to prime it with a small quantity of O.C.G. or a booster. It is a good practice to use high explosive primer
cartridges, at the top as well as at the bottom of the hole. Explosives are difficult to sink in water due to low
density of AN, and should preferably not be used in watery holes. If, however, AN fuel oil mixture has to be
used in watery holes, it should be packed in polythene bags and forced down the hole with the weight of a
high explosive and the stemming above. Holes of 62 mm dia. and above are considered economic for use of
AN-FO explosives. 8

The present day ANFO is unique, due to the use of porous prilled AN. The
properties important for explosive-grade prills are as follows:

1. Low clay content

2. Low moisture content
3. Free-flow sizing oil absorbency
4. Low particle density
5. Optimum friability
6. Non-caking tendency
7. Stability towards temperature cycling
 ▪ Reaction: On explosion, ANFO produces the following gases:

3𝑁𝐻4 𝑁𝑂3 + 𝐶𝐻2(𝑥) → 𝐶𝑂2 + 7𝐻2 𝑂 + 3𝑁2 + 𝐻𝑒𝑎𝑡

Ammonium Nitrate when reacts with aluminium produces the following gases:

3𝑁𝐻4 𝑁𝑂3 + 2𝐴𝑙 → 𝐴𝑙2 𝑂3 + 6𝐻2 𝑂 + 3𝑁2 + 𝐻𝑒𝑎𝑡

Previously ANFO used to be mixed & loaded
into the holes crudely by hand but presently automatic blending and loading machines are developed which
are widely being used worldwide. Fuel grade aluminium (particle size -20mes to +150mesh) nowadays adds
with ANFO to increase the energy and density of the explosive. About 15% aluminium appears to be the
economic limit from performance point of view. The major disadvantage of ANFO is that it is readily
desensitized by water and therefore can not be used in watery bore holes. If it is to be used in watery holes,
then it is to be packed in cartridge form, with the help of polythene packet. Nowadays ANFO incorporates
low density material such as polystyrene, sawdust, bagasse etc. for using in weak overburden materials and
saving costs. These low-cost additives can be added in percentage as high as 50% by volume to ANFO
without adversely affecting the fragmentation. For such mixtures, the typical detonation velocities are
around 2500m/s at explosive densities in the range of 0.4 to 0.5 gm/c.c. in 200mm. dia. Boreholes,
compared to 3500m/s for pure ANFO at a density of 0.829gm/c.c. The following precautionary measures are
to be taken while preparation & loading of the blastholes with ANFO:

1. Avoid Hand mixing of ammonium nitrate and fuel oil. Use special automatic blending machine made by
fire proof material to mix them inn a remote and isolated place.
2. The blending site should be free from any fire hazard.
3. Remove prepared ANFO as soon as possible to the blasting site by a vehicle.
4. Transport detonator, primer charge like OCG, special gelatine, etc. and detonating cord to the blasting
site by a separate vehicle.
5. Load ANFO into the blasthole with an automatic loading machine.

o Slurry Explosives: Since ANDO is hygroscopic in nature and has poor water-resistant property, the
development of AN based slurry explosives & Slurry Blasting Agent (SBA) have come into picture. The slurry
explosives are with Jelly like consistency and are water gels (The water-gel is a mixture of an oxidiser and
fuel sensitiser in an aqueous medium, thickened with a gum and gelled with a cross linking agent). In the
case of a permitted slurry, a coolant is added to reduce incendivity. The first commercial slurry explosive was
developed by Dr. Melville Cook in U.S.A. in 1957 consisted of TNT, AN and water in the ratio of 20: 65: 15. To
this, traces of chemicals for gelling and crosslinking were added to stabilise the homogeneity of the mixture.
In subsequent years the manufactures developed slurries with AN as the main ingredient and using variety
of sensitisers and fuels. The addition of metallic powder to the slurry enables the explosives to reach very
high strength. The chemical composition & property of Slurry Explosives are as follows:

S. Ingredient Percent Property Value

No.
1 Ammonium Nitrate 30-50 Density, g/cc 1.05-1.25
2 Sodium Nitrate 5-15 Detonation Velocity, 3800-4200
m/s
3 Water 10-25 Water resistance Good
4 Fuels 2-5 Fume Characteristics Excellent
5 Flake Aluminium 2-5
6 Mono-methyl Ammonium Nitrate 20-30
7 Guar Gum 1-3
8 Cross-Linking Agent 0.05-0.15
 Slurry explosives are among the safest of all explosives as
they are not easily ignited and are almost completely insensitive to the types of shock, bullet, impact and
friction stimuli that can initiate explosion of dynamites. In addition to the excellent safety characteristics of
these explosives, they have good fume characteristics, water resistance, etc. The slurry explosive has a sp. gr.
more than 1 and like ANFO can be poured directly into watery holes. They are also available in the form of
cartridges with plastic or polythene wrapper and some (permitted type) can be used in underground coal
mines.

The components required for ANFO and slurry explosives may be mixed at a plant
away from the blasting site or at the blasting site itself. In the case of PMS (plant mixed Murry) system, the
explosive is loaded into special tankers and from these tankers, the slurry is pumped directly into the blast-
hole. Where the volume of blasting is high enough to justify cost of transporting trucks, ICI has designed a
slurry pump truck which is capable of pumping Powerflow slurry directly into the blastholes. This system has
also achieved Slurry Automation. The ingredients are fed by automatically controlled pumps to a mixing
funnel, where they are picked up by a pump and forced rapidly through a long hose into the borehole, while
mixing & thickening are still in progress. Unlike cartridges slurries, pumpable slurries can be tailored to have
the appropriate density depending on strata conditions. In case of SMS system for blasting, one pump truck
can change nearly 25000 kgf in one shift. A small team of 5-6 professionally trained persons can load 50,000-
60,000 kgf of explosives into a large number of blast holes in a single working shift.

Slurry explosives will not explode accidentally dropped, or from shovel

impact, or even when involved in a fire. They have low non-toxic fumes and do not cause headache. Shelf life
of slurry explosives manufactured by most of the companies is usually one year. The best performance of
slurry explosives is obtained within four months from the date of manufacture.

o Emulsion Explosives: An emulsion is an intimate mixture of two liquids that do not dissolve in each other. In
more technical terms, an emulsion is described as a two-phase system in which an inner or dispersed phase
is distributed in an outer or continuous phase. Emulsions have for many years, contributed to our daily lives
in such products as Insecticides, photographic films and papers and cosmetics. The unique feature of an
emulsion explosive is that both the oxidizer and the fuel are liquids. The unique properties of emulsion
explosives are due to the minute size of the nitrate solution droplets and their tight compaction within the
continuous fuel phase. The basic formation of the emulsions is:

Ammonium Nitrate 60 to 70%

Calcium Nitrate/Sodium Nitrate 0 to 20 %

Fuel Oil 2 to 6%

Aluminium 1 to 3%

TNT and Water varies

The emulsion explosive is a mixture of oxidiser & fuel, both of which are in liquid
form. By an addition of emulsifiers, an intimate contact between oxidiser & fuel is obtained. The emulsion
slurry contains AN solution at high temperatures, combined with diesel oil, and an emulsifier which is passed
through a fast-moving blender. The blender mixes the ingredients into pumpable emulsion slurries with
grease like consistency. The resultant emulsion slurry is now made up of microscopic droplets of AN,
surrounded by fuel oil film and artificially created air bubble know as microspheres which makes the
emulsions detonable. There is no sensitisation by aluminium or explosive sensitiser but by mere air bubbles.
Thus, there is maximum energy generation because of intimate contact of oxidiser and fuel. In small
diameter products, glass micro-balloons and perlites are used to maintain sensitisation by aeration. There
are two types of emulsions:

1. Water in Oil Type

2. Oil in Water Type

Emulsion explosives depend entirely on the presence of

voids for initiation and propagation. A change in the amounts of voids effects a change in density. It
convenient and useful to relate properties to density and to consider voids and density adjusters. Slurry
explosives require thickeners and gelling agents to prevent segregation, to provide water resistance and to
control losses through cracks and fissures. Emulsion explosives cannot be gelled or cross linked. They do not
have the gel structure that characterises all slurry explosives. Velocity of detonation is a good Indicator of
reaction efficiency and is very dependent on particle size.

▪ Properties of Emulsion Explosives: are as follows:

1. High VOD: In view of the very fine particle size and intimate mix of oxidiser and fuel, the reaction
(detonation) through the explosive medium is very high. Emulsion Explosives have a velocity of
detonation between 4500-5500m/sec. This generates more shattering power.
2. Excellent Water resistance: Since every droplet of oxidiser is covered by a layer of fuel (oil/wax), the
water resistance of this explosive is excellent and comparatively superior to dynamites and slurries.
3. Safety: Emulsion Explosives are exceptionally safe during manufacture, transport, usage. They do
not contain any explosive ingredient. They are also comparatively insensitive to accidental initiation
by friction, stimuli & fire.
4. Satisfactory Low temperature detonatability.
5. Its velocity of detonation is high, and hence, a high detonation pressure is produced which gives
good fragmentation especially in hard rocks.
6. Because of the intimate mixture of composition, the efficiency is enhanced and unwanted fumes are
minimal.
7. They are less sensitive to accidental initiation through friction, static impact and fire. 5. They are
highly resistant to water.
8. They are not liable to pressure or shock desensitisation, a phenomenon known as dead pressing.
9. With on site preparation an explosive blend to suit the site can be formulated which results in cost
effective drilling and blasting.
10. Sensitivity to detonation can be varied.
11. Shelf life can be varied from a few days to several months or even years.

▪ Disadvantages: are as follows:

1. They are more fluid than slurry explosives and therefore create problems when loading a blast hole
with fissures or cracks.
2. They lack the strong cross-linked gel that characterizes the TNT-sensitized slurry explosives.

Emulsion slurries are claimed to have lower ingredient cost, higher density, higher
VOD and higher energy conversion and resultant bore-hole pressure as compared to water-gels/slurries.
Therefore, in low diameter sector, they are fast replacing slurries and water-gels. The emulsions are also
claimed to be sensitive even at smaller diameter holes, less than 25 mm. In addition, they are easily
pumpable without affecting their quality. Thus, emulsion slurries are hoped to be the low-cost, less
 hazardous replacement of NG based small diameter products. In emulsion explosives also, the commonly
used oxidisers are Ammonium Nitrate, Barium Nitrate, Calcium Nitrate etc. Fuel Oil, Mineral Oils, Waxes are
used as fuel.

o Heavy ANFO: The latest development of 1980's had been the use of emulsion slurries mixed with different
proportion of ANFO to give water resistant and higher density mixture which are named as Heavy ANFO or
HANFO. Thus, the emulsion to ANFO ratios can be from 20:80 to 50:50, depending on the severity of watery
conditions and need of stronger blast energy. The production of Heavy ANFO is done by mixing ANFO and
emulsion, which are stored in separate tanks on a bulk truck.

The emulsion occupies the interstices between the prills of ANFO within the
borehole The blends can be used efficiently in the larger volumes of blastholes, yield higher brisance, and
VOD, fit in any site-specific condition, yield better fragmentation and enhance movement of rock to suit
blast casting. The two constituents can be mixed in varying ratios depending upon sensitivity, energy, water
resistance requirements and economics.

One of the US manufacturers has successfully developed Heavy ANFO, using NGN slurry with
ANFO, Since the air spaces between AN prills are filled by emulsion, Heavy ANFO gives the advantage of
lower cost Like ANFOR with higher density, higher energy and better water resistance than ANFO or AN.
Thus, the mixture can have bulk density from 1.10 to 1.25 gm/cm. (compared to ANFO 0.8 g/cm), and bulk
strength almost 45% more than ANFO. The high energy ANFO through HANFO system or slurry concentrates
systems allows expansion of drilling pattern, thereby reducing drilling costs.

o Comparison between ANFO & Heavy ANFO: is as follows:

S. Characteristics ANFO Heavy ANFO Comments

No.
1 Density (gm/cc) 0.75-0.85 1.05-1.25
2 VOD (m/sec) 3000 3500-4500
3 Relative Weight Strength (ANFO-100) 1.0 1.2
4 Detonation Stability Fair Good Hence Heavy ANFO is
very good for
fractured Strata
5 Sensitivity Booster Sensitive. Booster Sensitive.
20-30% Primer made 20-25% Primer made
of higher strength of higher strength
explosive is required. explosive is required.
Cast booster cannot Cast Booster can be
be used used (0.2% of the
total weight of
explosives)
6 Coupling in the Borehole Very Good Excellent Better
Fragmentation is
achieved in Heavy
ANFO
7 Explosive Energy (kcal/kg) 900-920 1100-1150
8 Safety During Handling Safe Highly Safe Static Electricity may
cause an accident
9 Environmentally friendly Some Toxic Fumes Very Negligible toxic
are liberated fumes are liberated
10 Relative Cost related to cost of Higher then Lower Than
production of minerals Heavy-ANFO Bulk ANFO
Very less toxic fumes
Comparatively more sensitive (more than water)
Comparatively lesser cost than bulk Anfo
Greater VOD
Fits any site specific condition
Better Borehole Coupling
Yield higher brisance
Yield more fragmentation
o LOX: Liquid Oxygen liquifies at -163 degree Celsius. A given volume of liquid oxygen when gaseous, is
equivalent to 840 times at N.T.P., i.e., it has as much oxygen as would be available from 4000 times its
volume of atmospheric air. If a combustible ingredient, made in the form of a cartridge is soaked in liquid
oxygen and then subjected to reaction takes place with such terrific speed that large volume of a gas is
instantaneously released at high temperatures so as to cause explosion. The velocity of detonation under
suitable conditions of confinement can be faster than 5000m/sec. This is the principle behind its use as
explosive.

It is used for removal of overburden as well as mineral in the quarries of coal as well as
metalliferous mines. But its use is prohibited in underground coal mines. Like any other conventional
explosive, the Liquid Oxygen Explosive (LOX), comprises of two main broad constituents viz. the combustible
component and an oxidising agent. When the combustible component is saturated with liquid oxygen, each
particle of the combustible substance, is surrounded with an adequate amount of oxidising agent to ensure
complete combustion. Under conditions of confinement such as in a borehole, when ignited or detonated,
the reaction assumes explosive proportion resulting in the release of tremendous energy for blasting of
rock/minerals. LOX has a VOD in the range of 4000 to more than 5000m/s depending upon the liquid oxygen
content. These can be fired with or without help of detonator. Generally holes deeper than 3m. are fired by
a detonating fuse, as it is economical.

These cartridges are inflammable and the flow of gaseous oxygen emanating
from them will cause smouldering material, glowing coals, and cigarette stubs to burst into flames. It should
therefore be kept away from such burning or smouldering materials. Their characteristics are not constant
and depend upon the time elapsed between removal from soaking vessel and firing (typically 1 hr inn open
completely evaporates the absorbent). Hence, they should be fired without delay to prevent loss of
absorbed Liquid Oxygen. Grease or oil shall not come into contact with it at any stage. For use inn watery
holes, only H-type cartridges wrapped in polythene bag before lowering in the hole should be used. The toe
will not be sufficiently loosened, if LOX is used without such precautions in watery holes.

The explosives and performance characteristics of liquid

Oxygen explosive depend to a very large extent on the choice of the combustible substances which must
satisfy the following broad conditions:

1. The explosive made from it should be very safe during the carious operations of handling,
transportation, storage and usage.
2. These should have a high degree of inherent porosity so as to give a long and effective useful life.
3. These should easily mix with suitable additives to suit the varying strata conditions encountered in
different mines,
4. These should be easily available.

IOL’s Loxite explosive fulfils these requirements perfectly

and these are now being extensively used in the various mining industries such as iron ore, coal, bauxite,
magnesite etc. It is marketed by IOL, in cartridges of two types:

1. Small cartridges of dia. 25-900mm.

2. Large cartridges of dia. Over 100mm. up to 210mm.

For those areas, where the demand of LOX is heavy, IOL has established
central depots equipped with storage tanks and other arrangements of preparing it inn about an hour or two
before charging into the blastholes. Such depots are located at Bermo, Kathara, Lohardaga & in other
centrally located mining areas. These are transported in special vessels in trucks. For small consumers, within
300km from Oxygen factories, liquid oxygen is supplied in special containers by train, along with soaking
boxes & other equipment to prepare LOX cartridges on spot. Only small cartridges used for quarries are
prepared in this manner.

o Explosives used in Opencast Mines: Example of few explosives manufactured by IEL:

 S. Product Type Density VOD Water Diameter Cartridge Usage
No. (g/ml) (m/sec) Proofness (mm.) Weight
(kg)
1 OCG NG 1.4 6000 Excellent 75-200 2.5-12.5 As
Primer/Booster
& Toe Charge
2 Powerflo I Slurry 1.2 4000 Excellent 125-200 6.25-12.5 Column Charge
3 GN/1 AN Blasting 0.95 3000 Good 83-200 2.78-12.5 Column Charge
Agent
4 Powergel C Emulsion 1.2 5000 Excellent 83-200 2.78-12.5 Primer/Booster
5 Powergel 2 Emulsion 1.2 4800 Excellent 83-200 2.78-12.5 Base Cum
Column Charge
6 Powergel 1 Emulsion 1.2 4800 Excellent 83-200 2.78-12.5 Column Charge

o Comparison of Commercial Explosives: are as follows:

 Unit 3 (LO2)
o Initiation System: Commercial Explosives (and blasting agents) are designed to be relatively stable for safe
usage, transport, storage and manufacture. A powerful localised shock or detonation is required to initiate
commercial explosives. This is achieved by use of an initiating device, such as a detonator etc. An initiation
system consists of three basic parts:
1. An initial energy sources
2. An energy distribution network that conveys energy into the individual blastholes
3. An in-the hole component that uses the energy from the network to initiate a cap-sensitive explosive

The initial energy source may be electrical, such as a generator or

condenser- discharge blasting machine or a power line used to energise an electric blasting cap, or a heat
source such as a spark generator or a match. The energy conveyed to and into the individual blastholes may
be electricity, a burning fuse, a high-energy ex- plosive detonation or a low energy dust or gas detonation.

There are basically two methods of initiation, which are as follows:

1. Electrical Initiation Systems: These utilise an electrical power source with associated circuit wiring to
convey electrical energy to the detonators.
2. Non-Electric Initiation Systems: These utilise various types of chemical reactions ranging from
deflagration to detonation as a means of conveying the impulse to non-electric detonators, or as in the
case of detonating cord, it is the initiator. The first initiation systems t be used in mines, were non-
electric type only, but were gradually replaces by the Electrical ones, due to certain advantages.
 However, electrical firing involves problems of detonators and extraneous electricity. In addition, there
are severe demands on the exploders. Several non-electric firing systems have become popular in the
last few years.
o Shock Tube Initiation System: In normal deep-hole blasting, the detonating fuse, de-sensitises the explosive
of the hole to some extent. It also disturbs the stemming of the hole. This reduces the efficiency of the
explosives. In order to remove the issue, Shock Tube System is employed. This is a system of non-electrical
firing, that does not utilise detonating cord. In this system, a polymer tube (external dia. 3mm and Internal
dia. 1.5mm.) is used in place of Detonating Fuse. One of its end is connected with a ms delay detonator, in
the production itself. This end is lowered in the hole, and the other end is connected has a plastic connector,
that is connected to the trunk detonating Fuse on the surface.

Inside these tubes is an explosive material that propagates a mild detonation which
activates the cap. On being initiated by the electric detonator on the surface, a shock wave travels inside the
shock tube, with a velocity of around 2000ms. This wave fires the detonator of the tube, thus exploding the
charge, and during this process, the shock tube remains undamaged. These initiation systems are not
susceptible to extraneous electricity, create little or no air-blast, do not disrupt the charge in the blastholes,
and have delay accuracy similar to those of electric cap systems. Several systems are under development or
in completion stage. These systems are being used in soke opencast mines in India, where controlled
blasting is required near the residential areas. Despite being costly, it is advantageous as it reduces the
ground vibrations and it also increases the explosive efficiency. It also reduces the air-blast and noise greatly.

The Shock Tube Initiation System consists of three sub-systems, which are as follows:

1. Nonel: Nonel stands for Non-Electric Detonator. It is the common trade name of a series of blast
initiation accessories, developed by Nobel AB of Sweden. It uses the Shock-Tube principle. It is a safe,
multipurpose firing system which combines the simplicity of fuse and igniter cord with the precision of
the electric firing.

It consists of a flexible plastic tube having 3 mm. external and 1.5 mm. internal
diameter. The tubes are available in pre-cut lengths. Its inside is coated with an explosive composition
having a detonating velocity of 1800m/s. This shockwave has sufficient energy to initiate the primary
explosive or delay element in a detonator. Since the reaction leaves the shock tube intact, there is no
lateral shock effect and the tube acts merely as a signal conductor.
Primers of explosives with Nonel detonators inserted in them are charged in
the blast-holes and the Nonel tubes are bunched for convenience of connection to the mains blasting
system. Upon initiation, the shock wave passes down the plastic tubes, the insides of which are coated
with reactive substance that maintains the shock wave at a rate of approx. 2000 m. per second.
One end of the tube is fitted with a non-electric delay detonator
which is crimped to it in the factory while the other end is sealed. The end having detonator is lowered
down into the blast hole while the sealed end projects outside the hole. The sealed end is initiated by a
detonator or detonating cord. Nonel without a cap can not detonate on its own, any commercial
product. The shock wave passing through the tube is very mild and cannot even damage the walls of the
 tube as the quantity of explosive in the tube is very low. Nonel or Shock tube system has two good
characteristics compared to the detonating cord which are mentioned below:
a. Noiseless Characteristics
b. It has no effects on the explosive column.
These can easily be used in external disturbances like electric storms, static electricity, stary
currents etc. and since each compound in the system is sealed in the factory, it can be nicely used in
underwater conditions. It is also immune to misfires caused by current leakage in conductive orebodies
and eliminates the need for complicated electrical circuit testing and shot-firing equipment. Raydet
manufactured by IDL Chemicals, is just like Nonel. Some of the latest models of the Nonel Series are
Nonel GT-HD, Nonel Gt-OD, Nonel Unidet, Nonel GT-MS, Nonel GT-T etc. The nonel system can be
initiated by an ordinary detonator or an electric detonator or special starter caps.

2. Exel: The shock tubes manufactured by American manufacturer (Orica-IEL) is marketed under the brand
name of Exel. The original shock tube of the late 1970s was composed of an ionomer resin which was
well-suited for its purpose on several accounts but deficient in others; namely strength, toughness and
resistance to contamination by water or diesel fuel. Current shock tubes are an improved type
introduced in the early 1980s and are overcoated with polythene. However, with improper loading of
heavy cartridges in deep holes, even this outer layer of polyethylene may not prevent stressing of the
tube beyond its breaking point resulting in a hole cut-off. A new type of shock tube has been introduced
called Exel which consists of a plastic tube with an outside diameter of approximately three mm., the
inside walls of which are coated with a fine layer of explosive dust. The dust is typically a mixture of HMX
(Octogen) with a small percentage of flake aluminium. Exel tubing is able to withstand rough handling, -
low temperatures up to -3.5 degrees Celsius & high surface temperatures of 65 degrees Celsius and is
impermeable to diesel oil used in ANFO-type explosives and water/ammonium nitrate solutions in wet
holes.

3. Raydet: This pproduct has been developed by IDL Industries Ltd. & is very similar to Nonel using the
Shock Tube principle. Raydet is a non-electric initiating device combing the versatility and advantages of
electric delay detonator and detonating cord. It also overcomes some of their disadvantages. Raydet can
also be used for Surface Blasting, Quarrying, Trenching, Large Hole Stoping, Ring Hole Blasting,
Tunnelling and other underground blasting operations.

It consists of a plastic tube, having an internal diameter of 1.5mm. &

external diameter of 3.0mm., carrying a very small quantity of explosive material on its inner surface
(around 20mg/m). A high strength no. 8 instantaneous or delay detonator is crimped to one end of the
ray tube. The other end is sealed to prevent ingress of moisture and other foreign matter. When
initiated, a low order shock wave travels through the tube and reliably initiates the detonators. Raydet
can be initiated by a detonator or a detonating cord. A tag indicates the delay number of raydet and a
tape fastening the tube in a coil indicates the tube length. The detonation velocity passing through the
shock tube is of the order of 2000m/s. In view of the low charge inside the tube, even after the
detonation is complete, the tube remains intact. Hence, unlike detonating cord, this does not disrupt the
explosive columne in the borehole.

20 Delay periods in the millisecond range and 12 periods in the half second
range are currently available. The delays are from no. 0 delay to no. 15 delay; No. 0 delay is
instantaneous No. 1 delay is 50 ms and No. 15 delay is 625 ms. These are available with different tube
lengths to suit customer’s requirements.

When using the raydet, do not cut factory sealed end of ray tube and do not connect two ray
tubes. One ray tube will not initiate another. Plastic connectors specially designed for hook-up of Raydet
tube onto detonating cord are provided with every Raydet. Bunch Blasting up to 10 ‘Raydet Tubes’ using
a wrap of ‘Detonating Cord’ is also possible especially in tunnel blasting, ring blasting, stoping etc., which
makes hook-up very easy and less time consuming.

▪ Advantages: the advantages of using Raydets are as follows:

1. Raydet being a non-electric system, is immune to extraneous electricity sources such as stray
currents, static electricity and radio frequency energies emanating from walkie-talkies.
2. Being Non-Electric, it eliminates totally the risk of misfires of electric detonators due to earth
leakage in highly conducive ore-bodies
3. Raydet Tubes are insensitive to accidental initiation due to fire, impact and friction normally
encountered during handling/usage. Detonators used are conventional ones and are required to be
handled like other detonators.
4. Raydet provides accurate down-the-hole delay and in surface blasting, provides true bottom
priming. Bottom Hole initiation is a pre-requisite for efficient bench blasting, results and reduces the
risk of fly rock generation, formation of toe and air blast noise.
5. Upon determination, ’Detonating Cord’ generates a lot of shock and gas. In boreholes, this results in
ejection of stemming material resulting in premature venting of detonation products and also partial
de-sensitisation of non-cap sensitive explosives in the borehole. Detonating Cords of higher
grammage can also cause, at time, detonation of explosives across the borehole diameter (instead of
along the column). The explosive thus initiated gets consumed before steady state VOD is reached
and thus results in less energy release.
It does not cause any disturbance to the stemming and also ensures full energy
release from the explosive present in the borehole. Raydet tube would not initiate any cap-sensitive
explosive in the borehole unlike other delay systems based on low energy detonating cord (LEDC).
6. Where needed, Raydet down-the-hole in conjunction with surface delays, provides a large number
of sequential delays. Multiple delays enhance fragmentation, well spread-out muck-pile and
minimises ground vibrations through reduced maximum charge per delay.
7. Initiation of explosive decks on different delays in the same borehole is possible, such as for
controlling ground vibrations in surface blasting and in large hole Stoping in the underground.
8. 'Bunch Blasting' of up to 10 Raydet tubes with a single wrap of detonating cord is possible in
tunnelling and stoping operations, which makes connections easy, less cumbersome and time
saving. In bunch blasting, use of plastic connector is eliminated.
9. Raydet does not cause any desensitisation of booster-sensitive explosive in the blasthole. Hence full
energy of the explosive can be realised in the blasthole.
10. It does not cause any cratering effects from the collar portion of the blasthole.
11. Explosive decks in the same blasthole can be initiated on different delay intervals to reduce
maximum charge per delay.
 12. Since the detonation wave is contained within the tube, it significantly reduces air-blast noise when
used on the surface in place of detonating cord trunklines.

Raydet initiation system resulted in considerable

improvement in production, productivity, cost reduction, better fragmentation, higher loading
machine output, and presented itself as a better environmentally friendly system.

▪ Raydet Products and Uses: are as follows:

1. Raydet HS (Half second) Delay no. 1-12.
Uses: Tunnelling. Shaft sinking, etc.
2. Raydet MS (Milli Second) Delay No. 1-20.
Uses: Tunnelling, shaft sinking, underground stoping, ring hole blasting, long hole stoping, drop raising,
blasting in opencast mines, trenching, etc.
3. Raydet DTH (Down the hole) 475 and 500 ms.
Uses: Long hole stoping, drop raising, vertical crater retreat (VCR) blasting, pipeline trenching, etc.
4. Raydet TLD (Trunk line delay), 0, 17, 15, 42 and 67 ms
Uses: In open cast mines and trench blasting as surface delay
5. TLD 'o' ms delay can be used for 'Bunch Blasting' of Raydets (ms/HS) in tunnelling/excavation for
hydroelectric, rail and road projects, shaft sinking and also in the underground metalliferous mines.

o Electronic Detonator: These are based on Micro-Chip Technology & use pyrotechnic delay elements. These
are modern, fully programmable and intelligent high strength detonators. These are basically a device, which
stores electrical energy for a certain time and then deliver that energy as a sharp pulse at a precise time to a
conventional blasting cap or electric blasting circuit. With these types of delay detonators, possibility of
overlap is completely eliminated. Static Electricity, radio Frequency or Stray Currents have no influence on it.
The Electronic Delay Detonator Firing System comprises the following components:
1. Up to 200 integrated electronic detonator firing caps,
2. A console for the testing and programming of each integrated electronic detonator (IED).
3. A bifilar line which links the firing console to the firing stand,
4. A firing console for operational check and the management of the firing itself.

Each detonator consist a circuit board that can be programmed to initiate at precise
millisecond timing with a firing sequence. Each detonator has a unique detonator ID number allotted to it at
the time of manufacturing and are individually programmed via the use of a handheld tagger and initiated
via a bespoke blasting box. The delay element is a capacitor controlled by application-specific integrated
circuit (ASIC) fitted before igniter and the base charge. Electronic blasting machine is the only devices that
are designed to provide password protection, programming capability as well as the energy levels needed to
charge the detonators in a circuit and send a fire command.
 Except incorporating a slightly lengthened version of the current PVC plug the
detonator looks identical to an instantaneous detonator.

▪ Advantages: The Electronic Delay Detonators Offer the following Advantages:

1. Precise Delay: Since delay is achieved electronically , these are very precise.
2. Programmable Delay: Delay interval can be programmed at site as per the requirement.
3. Safe to use in place where there is a risk of stray current.
4. Only authorised person having the key can fire the blast.
5. offer flexibility and reliable sequencing.
6. provide accurate timing, reduce ground vibration, improve fragmentation, enhance selectivity,
reduce blast-oriented damage,
7. reduce operating cost,
8. improve delay precision,
9. offer re-programming of delays of detonators at any time till the moment before firing.
10. offer automatic checking facilities, offer facilities of easy, quick and most accurate way of
preparing a blast for firing a complex circuit,
11. reduce over-braking problem, offer much safety in initiation process.
12. Offer much safety in initiation process.

▪ Disadvantages: are as follows:

13. Costly as compared to other initiation systems.

---------------------------------------------------------Extras---------------------------------------------

o Detonator: A detonator is a small auxiliary charge of special explosive enclosed in a small copper or
aluminium tube. A chemical reaction initiated, either by a flame (safety fuse) or electric current in the
special explosive can build up very rapidly, and the special explosive explodes with high intensity and
propagates detonation wave through the high explosive around its external periphery & initiates the latter.
Detonators are of following types:
1. Plain Detonators: These are fired by safety fuses, the spark or “spit” from the fuse causing the detonator
to explode; these are sometimes called “Ordinary” detonators. These consists of a base charge of
pentaerythritol tetra nitrate (PETN) with a priming charge of ASA composition (lead azide, lead
styphnate and aluminium powder) in an aluminium or copper tube. The ASA initiates the base charge. It
is insensitive to shock, friction and lightning.
2. Electric detonators: These are fired by passage of electric current through the detonators. They are
subdivided as:
 a. Low Tension Detonators
b. High tension Detonators
They are of copper or aluminium tubes. These are fired by passing current through a
fuse-head. A current of 0.5amp at a minimum 3.5 volts can ignite the fuse-head composition which in
turn initiates the priming charge. These are essentially of two types on the basis of time of action:
1. Instantaneous Detonators
2. Delay Detonators

o Bulk Explosive System: These are essentially those where explosives are delivered directly into the borehole
through mechanised and mobile delivery systems. Charging Rates are fairly high and capacities of a single
delivery unit ranges from 7-25t per load. In view of these features Bulk Loading Systems are ideally suited to
large explosive consumers with consumption levels of 2000tpa or more although lower consumption levels
may also be serviced at a higher cost. If the density of the bulk explosive goes beyond 1.3gm/cc, the
detonation may not be reliable and results in poor VOD, incomplete detonation & poor energy release.
Products with low density are not selected since they have low bulk strength and also cause trouble in water
filled hole. A product with a density ranging from 1gm/cc to 125gm/cc is ideal, which will offer best
detonation characteristics and energy release.

With ideal consumption levels, Bulk explosives offer the following advantages:

1. Safety: Non-Explosives ingredients are stored at all stages

2. Inventory: No inventory needs to be maintained and hence no investments in large magazines are
necessary
3. Explosive vans: Large Number of vans are not required
4. Manpower: Manpower savings are obtained with less deployment of van drivers/helper, blasting crew
and magazine staff.
5. Speed of Operation: Swift charging rate of around 250-300 kg/min.
6. Explosive Product: Tailor made product with respect to density and energy can be delivered down the
hole
7. Blasting Efficiency: Full Borehole coupling can be achieved enabling expanded burden/spacing
parameters
8. Other Features: No explosive pilferage
9. Supply Source & Magazine: Nearly supply source and smaller magazine is required which will reduce the
transportation cost, faster transport of explosives and require cheaper magazine.
10. Operational Satisfaction & Bigger Blast: Because of clean, less complicated methodology, safer
technique, easiness to handle and operate, the blaster always becomes very happy with the system. It
also offers to blast a huge number of blastholes at a time which definitely reduces the frequency of blast
per weak & enhances higher output/blast, production time, production & productivity.
11. Use of Larger Diameter Blastholes: With this system large diameter blastholes can be efficiently be
used, which will decrease drilling costs per tonne of output.
12. Better Fragmentation: With best blast design, better fragmentation will be obtained. This will enhance
the production and productivity of dumpers, shovels/draglines/payloaders, crushers etc. with lower
power costs, higher utilization of machine, higher bucket fill factor of shovel/dragline etc
13. Environmentally friendly: It is environmentally friendly as very low quantity toxic fumes are produced.

Apart from above positive features, bulk plants require fixed

investments and with higher offtake levels, the cost gets distributed over a wider volume enabling explosive
manufacturers to offer lower prices, Hence, cost savings are even greater at larger offtake levels. Explosives
amenable to bulk loading are ANFO, slurries, emulsions/doped emulsions and emulsified Heavy ANFO. With
the availability of a number of bulk explosive systems, the choice is a difficult one. Several factors need to be
considered:

1. Normally for large operations bulk loading systems are suitable. Problems may arise when one comes
across limited size of individual mines and quarries. For smaller operations the shared services of a pump
truck and its support facilities may be necessary. In such situations to operate this system effectively a
strictly prearranged schedule of hole loading may be necessary
2. Not only blasting must be arranged to a given schedule, but also the size of the shots be tailored to suit
the capacities of pump trucks and support facility.
3. Rock and blasting conditions have very important effect. For soft to medium strength rock such as in coal
measure strata most of the systems work satisfactorily. But in hard rock it seems emulsions work better.
In watery holes ANFO usage is precluded and even with Heavy ANFO (with up to 25% emulsion) reliable
blast results are not achieved. When hard rock and watery holes are encountered emulsions are found
to be much better.
4. Whenever, any mine is planning for a bulk loaded explosive system it must be considered that variable
rock conditions will be encountered hence any system. which offers more than one type of explosive
with variable composition, be chosen. This is an essential consideration since mine will then be
dependent on one plant only, hence that plant should be able to supply all the required types and
compositions of explosives:
5. An important advantage of bulk loading system is that explosive companies are responsible for carrying
out charging and assisting in blasting operations and hence they provide technical service. A company
with good technical background is an obvious choice.
6. The system should have a check on the weight and quality of the explosive delivered. It is difficult to
have full-proof system hence, it is advisable that double checking system including electronic records are
adopted.
7. The system should have adequate alternative possibilities so that impact of breakdown of the system is
reduced to the minimum.
8. Another relevant consideration is about the usage of increased explosive charge per hole (up to 20%)
because of the better coupling afforded by these types of explosives. This may pose problems if the mine
blasting is near an urban area.

There are at present, 4 type of delivery systems mainly inn use & these are:

1. Plant Mixed Slurry/Emulsion

2. Site Mixed Slurry/Emulsion
3. Re-pumpable slurry/emulsion
 4. ANFO/HANFO delivery system

o Site Mixed Slurry or Emulsion System (SMS/SME): These systems consist of support plant and pump trucks.
The support plant is located near the mine or at centrally located place if it caters to a group of mines. Non-
Explosive ingredients are stored in it. Certain intermediates are prepared from some of these ingredients
and kept ready. When the blastholes are to be loaded, the ingredients are loaded in specially designed pump
trucks.
The pump trucks are specially designed to carry all the ingredients and to pump the
blended slurry into boreholes through a delivery hose carried on the truck. The blending operation is
controlled by a sophisticated control system. A predetermined quantity can be pumped into a hole through
the control system. The various ingredients are continuously metered and passed through a hose into the
borehole. In case high energy bottom load and low energy top load is required, the desired quantity of each
is set on the control system. At the site, the calibrated quantities of ingredients for a particular product are
mixed and the product is pumped into the blast hole through a delivery hose. The mix becomes sensitive
after five minutes into the hole after certain reactions are completed. In this system there is a flexibility to
charge three different product formulations of various energies and densities in predetermined quantities in
the same hole.
The pump truck is a vital part of SMS system. These trucks are specially designed to
carry all the ingredients required for blasting and to pump the blended slurry into boreholes through a
delivery hose carried on the truck. At the blasting site the above ingredients along with a suitable cross-
linking agent like guar gum (to form gel) and gassing agent (for controlling density of slurry) are all mixed
together just before commencing the loading itself so as to form a pumpable slurry of the required com-
position. The blending operation is controlled by a sophisticated control system which is reliable and simple
in operation. A pre-determined quantity to be pumped into a hole can be set through the control system by
the operator and then a 'start button' is pressed to pump the required slurry into the borehole. The various
ingredients are continuously metered and passed through a hose into the borehole. When the quantity of
slurry is delivered the unit automatically shuts off. In case high energy bottom load & low energy top load is
required, the desire quantity of each is set on the control system. After the desired quantity of bottom load
is delivered, the unit automatically shifts to the low energy composition. When the desired quantity of both
bottom and to loads are delivered, the unit automatically shuts off. This ensures that there is neither excess
explosive getting into the borehole, nor stores in the pump trucks. There is a facility of charging three types
of explosives having varied strengths into the blast holes depending on the nature of strata and sequence as
desired. The special precautions to be taken in SMS/SME system, are as follows;
1. Since the explosive is not inn cartridge form, permission shall be taken from the DGMS.
2. All blasting operations shall be carried out under the personal supervision of an undermanager and an
engineer of the explosive supplying company.
3. Smoking, welding or any other open light shall not be allowed within 60m. while charging. The ground
up to 10m. from the pump truck shall be free from dry grass.
4. Pump truck shall be kept earthed by chain. The hose-pipe shall be of anti-static type..
5. Only minimum number of persons shall be present while charging.
6. For sleeping holes, and for blasting in hot strata permission shall be obtained from DGMS.
7. Primer explosive cartridge shall not be slit or deformed.
8. A sketch of the site showing location of holes, type of strata, burden, spacing, depth, hole dia, quantity
charged in each hole, length of stemming, type of explosive, total quantity of explosive used, date and
location of blasting shall be prepared by under manager incharge of blasting. A record of results of
blasting (fragmentation, flyrock distance, vibration, back break and oversize) shall also be kept.
9. Mine staff engaged shall be given proper training.
10. The supplier shall get the pump truck approved by the Controller of Explosives once in a year.
 o Comparison Between Cartridge & Bulk Explosives: The comparison between the two systems, for a mine
using about 5000 tonnes of explosives per year, is as follows:

S. Criterions Cartridges Bulk Explosives

No.
1 Safety Explosive is handled, stored and transported. Explosive handling, storage 7 transportation
Thus, potential hazard is always there. is totally eliminated. The ingredients become
explosive on being charged in the hole.
2 Explosive For the consumption level of Storage facility for only cast boosters
Magazine 5000tonnes/year, at least 200 tonnes storage needed. Cast Boosters used is 0.1% or
(Size) magazine are needed. 500kg/year. Thus, 500kg magazine is
adequate for cat booster and small capacity
for IDF & detonators.
(Cost) The cost of magazine of 200 tonnes would be The cost of magazine would be small.
around 20 times.

(Safety Zone) Large Safety Zone is required which would be A very small safety zone of 95m. radius is
difficult to get. The safety zone for 4(50t) needed.
magazines has to be 820m radius with 90m
distance between magazines. Thus, for 4
magazines, a safety zone of 1km. radius is
needed.
3 Explosive For charging 400tonne if explosives per A small explosive van of 500kg capacity
Vans month at least 3 vans of 10 tonnes capacity would be sufficient.
(Requirement) are needed for transportation of explosives
from magazine to site.
(Cost) The capital cost of 3 vans is nearly ten times. The capital cost of these vans would be
small.

(Manpower) Three van drivers & 3 helpers would need to One driver & one helper would be needed.
be employed.

4 Blasting Crew The Blasting Crew comprising of at least 12 The blast crew would comprise of 2-3
labourers and two supervisors would be workers & one supervisor.
needed for charging & stemming.
5 Charging Rate For charging/Stemming of 20 tonnes of For charging 20 tonnes about 3 hours are
explosives, a full shift would be needed using needed using 2-3 workers for stemming. The
12 labourers. operation is fully mechanised.
6 Explosive Explosive of fixed energy is available. The A wide range of explosive energies are
quality available in stock has to be used. available. There are over 20 formulations to
Inventory of each product is required to be choose form. These can be made with minor
maintained adjustments on pump truck itself.
7 Explosive The density of the product cannot be varied The density of the product can be varied to
suit the requirements over a wide range.
8 Coupling The coupling is not full. The coupling is full resulting in maximum
transfer of energy from explosive to rock.
9 Cost The cost is fixed The cost decreases as offtake increases.
10 Pilferage There is hazard of explosive pilferage. Pilferage is not possible.
o Precautions for bulk transport- are as follows:
1. The explosive van shall be got approved by the Controller of Explosives
2. Permission shall be taken from the Regional Inspector.
3. Explosives and detonators shall not be loaded together in the vehicle and the quantity of explosives/
detonators transported shall not exceed the approved quantity. The explosives cases shall not be piled
higher than the sides of the body.
4. The vehicle shall be marked on both sides and ends with the word 'EXPLOSIVE' in red letters not less
than 25cm high on a white background.
5. Vehicle shall be provided with 2 fire extinguishers (one of Carbon Tetra Chloride type for petroleum fire
and the other of CO₂ pressure type for electrical fire) suitably placed for immediate use.
6.
a. No person other than the driver and his helper shall ride in the vehicle.
b. Vehicle loaded with explosives shall not be left unattended.
c. The engine shall be stopped and brakes applied securely before it is loaded or unloaded or left
standing.
d. Speed shall not exceed 25km/hr The vehicle shall not be taken into garage or repair shop and
shall not be parked in a congested place.
e. The vehicle shall not be re-fuelled except in emergency; during re-fuelling, the engine shall be
stopped and other precautions taken to prevent accident.
7. All operations connected with transport of explosives shall be conducted under the personal supervision
of the blasting foreman.
8. The blasting foreman shall make sure that no person engaged in loading and transport of explosives has
cigar, cigarette, bidi, or match-stick etc in his possession.
9. The explosive van shall be carefully inspected by a competent person once in 24 hours. Fire
extinguishers, insulation of wires, cleanliness of vehicle, leakage of fuel, lights, brakes and steering shall
all be checked. Record of the inspection shall be kept in a bound paged book.

Unit 3 (LO3)
o Different parameters connected to Drilling of Blast Holes: are as follows:
1. Burden: Burden is the distance of hole from the free face or from the nearest hole in the next row. It is
an important aspect in the blasting practice. It is based on the following:
a. hardness of the strata
b. hole diameter
c. explosive strength and density
d. bench height
e. bucket size of loading machine.

On the basis of hardness of strata, the typical values of Burden for different
strata, is as follows:

S. Hardness of Strata Strength of Explosives

No. Medium High
1 Soft Rocks/Coal B=35d B=40d
2 Sand Stone B=30d B=35d
3 Hard Sand Stones B=25d B=30d
4 Iron Stone/Granite B=20d B=25d

Where, B = Burden

d = diameter of hole.
 Generally, the burden is reduced in the third and their
successive rows (0.9B), and if the face has a Toe, then the Burden is reduced. In the case of strata with
geological disturbances, such as Joints, etc., the orientation & the condition of the joints is also taken into
consideration, while determining the Burden. If a joint is parallel to the face, then the burden is kept less.
When the Hole diameter is less and the burden-spacing is also less, then the material is broken into
smaller pieces but the cost of drilling & blasting is more. On the other hand, when the hole diameter and
the Burden-Spacing is more, the cost of Drilling & Blasting is less.

If the burden is very less:

▪ The material will be thrown very far away from the face
▪ There will be problems of air-blast and fly-rocks
▪ The material will be crushed

If the burden is greater than required, then:

▪ The fragmentation will be poor

▪ Toe will be left behind.
▪ There will be problem of back-break and quarry side will be damaged.

There are two formulas for the

estimation of Burden:

▪ Konya & Walter (1990):

2𝑆𝐺𝑒
𝐵=( + 1.5) × 𝐷𝑒
𝑆𝐺𝑟
Where, B = Burden (ft)
SGe = Specific Gravity of Explosive
SGe = Specific Gravity of Rock
De = Diameter of Explosive (Inch)

▪ Pal Roy (2005)

𝐷𝑒 𝐿
𝐵=𝐻× + 0.37 ( ) 0.5
𝐷ℎ 𝐶
Where, B = Burden (m)
H = Height of Bench (m)
De = Diameter of Hole (mm)
Dh = Hole Diameter (mm)
RQD = Rock Quality Designation
L = Loading Density of Explosive (kg/m)
C = Charge Factor if Explosive (kg/m3)
2. Drilling Method: Rotary & down the hole drills are commonly used for drilling in Indian opencast mines.
DTH are compressed air operated at a pressure of 7kg/cm2 with hole diameter 100mm. to 225mm. and
generally used up to a depth of 12 to 15m. However, DTH is operated at a pressure of 14.1kg/cm2 with
hole diameters of 311mm. to 380mm. may be widely used inn Indian Coal Mines in future. Rotary drilling
method used in India for large diameter hole & in rocks of very high compressive strength. Generally,
tricone roller bits in conjunction with high pull-down force are employed for satisfactory penetration
rate. Drills up to 400mm. dia. And a depth of more than 20m. can be practised with the rotary drilling
method in hard rock formation.

3. Spacing: Spacing is the distance between two holes in a row. It is majorly depended upon the Burden.
▪ Generally, Spacing = (1.2 – 1.5) B.
▪ If the hole diameter is large, the spacing is kept around 1.2-1.3B, and if the hole diameter is small,
the spacing is kept around1.5B.
▪ Spacing is kept 2B, if there is no delay from hole to hole.
▪ In the case of Cast Blasting, the spacing is kept 1B or less.
▪ It also based on the hardness of the strata:
Hard Stone – Burden * 1
Medium Stone – Burden * 1.25
Soft Strata/Coal – Burden * 1.50

Taking the above factors into consideration, the spacing and Burden is
calculated, and then Trial Blasting is done. If the Burden is increased inn place of Spacing, the material
forms bigger Boulders. Hence, whenever it is required to expand the pattern, in order to save on
explosives, the trial shall be done by increasing the spacing in place of Burden. If the spacing is more
than required, a “Hump” is left on the base of the bench.

4. Drilling Pattern: This may be:

a. Grid Pattern: In this, the holes of different rows are exactly one behind the other. It is used for
coal and soft strata.
b. Staggered Pattern: In this, each hole of a row is between 2 holes of an earlier or later row. This
pattern is used for medium & hard strata.

5. Sub-Grade: In very Hard Strata, some material falls into the hole while drilling o while removing the drill
rod from the hole. In such cases, the material is not fully broken and a toe is left on the bottom. In order
to combat this problem, hole is drilled to some distance below the floor of the bench so that the toe is
not left. This distance is known as sub-grade. Usually,
Sub-Grade = Bench height * 0.1

Sub Grade drilling is done only in rocks, and not OB

bench or coal. This is because, if it is done, the coal will be wasted and it will be problematic to operate
heavy shovels in it. Unrequired sub grade drilling leads to wastage of explosives and increased ground
vibrations.
 6. Position of Booster: Usually, the booster is placed at the bottom of the hole, but where there is water or
muck in the hole, it is hung 1 to 1.5m. above the bottom of the hole.

7. Explosive: Volume of material is calculated as burden * spacing * hole depth. Explosive Charge is
decided as follows:
Hard Stone – 0.4kg/m3
Medium hard Stone – 0.3kg/m3
Soft Stone (shale)/Coal – 0.2kg/m3

If there are hard & soft strata in the same bench, Deck
Charging is done. In blasting by SMS, specific gravity of the explosive can be varied according to
requirement. Charge per m3 is decided after a few trials.

8. Stemming Height: Stemming prevents the escape of gases produced during the blast and keeps them
confined. A proper stemming prevents fly-rocks, provides good fragmentation and proper muck-pile. The
height of stemming is called Stemming Height. It should be equal to or more than the burden distance.
Inadequate stemming causes fly-rocks and the material does not fragment well.
(0.7 to 1.0) B

If the rocks overlying the hole are hard, the stemming height is
kept 0.7B, if soft stone is present, and there are natural cracks or water in the strata, then the stemming
height is kept 1.0B. If the stemming height is less than required, there is more possibility of Fly-rocks
and blasting noise. If it is more than required, then the size of material from the Stemming Zone would
be more, and boulder will be formed near the mouth of the hole.

9. Delay Interval: A rough rule is:

Hard Stone – Burden (in feet) * 3ms
Medium Hard Stone – Burden (in feet) * 4ms
Soft Stone/Coal – Burden (in feet) * 5ms

10. Initiation Pattern: Hole connection or Initiation Pattern is of two types:

a. In-Line Pattern: where the holes are blasted line by line. This pattern is preferred for soft stone
and coal.
b. V pattern: where holes of one row are blasted with some holes of the later rows. This pattern is
good for hard stone.

A detailed record of blasting parameters and the results obtained is

maintained, and based on experience, the parameters are varied to arrive at an optimum blasting
pattern.

11. Bench height: In opencast coal mines majority of bench height is 6 to 12m. where 4.6m3 or 6 m3 shovels
or 8 m3 front end loaders are in use and with 10m3 capacity shovel bench height is generally 10-12m. In
some of the big opencast mines where 24 m3 bucket capacity dragline with 96m. boom are in use, there
bench height is more than 30m.

12. Blast hole Diameter: In opencast coal mines, the diameter of blasthole is generally 100, 150 or 163 mm.
up to a bench height of 8m. bench height 250mm. blast hole diameter is chosen.

o Patterns of Drill Holes:

▪ Advantages: During drill pattern local geology, location of free faces, shape of the area,
fragmentation planned, dimension of boreholes and explosives, depth of holes, local blasting
experience, etc. play an important role. An Improper drill pattern will lead to blown-outs, fly-rocks,
air-blast, heavy ground vibrations, poor fragmentation, over break or under break, toe formation,
wasting of explosive energy, strains on the loading machine, secondary blasting and overall
uneconomic. Basic drill patterns are:
1. In-Line Square or Rectangular Patterns: In this type, the holes of every row are exactly in line,
with the corresponding hole in the next row. Inn this way, a grid pattern is formed If the burden
& spacing are equal, then the square pattern is formed, and if the spacing is greater than the
burden, then the Rectangular pattern is formed,

2. Staggered: In this pattern, a hole of any one row, falls in between any two holes of the next row.
Generally, the Staggered pattern results in better blasting, especially where an equilateral
triangle is formed.

In case of a block, where two free faces are available, a square

pattern gives better results. The staggered pattern is better for line-by-line firing, and the square
or rectangular pattern is good for V cut firing. Other patterns are chosen based on the
manipulation of these patterns.

3. Single Row Firing: In single row firing, the choice between firing he holes simultaneously or in a
sequence depends on the nature of the rock and the degree of fragmentation and throw
desired. Delay firing of the series of holes in a row is done to obtain better fragmentation. The
other advantages of delay firing are the reduction in ground vibrations , less back-break or over-
break, and better control of the rock pile.

4. Multi Row Firing: In This, the two basic short delay firing for single row can be applied in various
ways, as shown in Fig. D,E,F,G. A few of them have been illustrated. In initiating multi—row
rounds of holes, two factors are important, viz.,:

a. The point where the breakage begins on the first row as shown by ‘a’ in fig. D should as
far as practicable be located where either the rock structure is weakest or the burden is
most favourable, and,
b. The inherent ignition scatter of detonators of the same delay period should be taken
into account while designing the ignition patterns that every hole has a free breakage
zone in front as illustrated by fig E.
▪ Patterns of Drilling: For bench blasting the shot-holes may be drilled in two ways. The choice of the
drilling pattern depends on the conditions prevailing in the mine as given below:
1. Vertical Holes: The vertical holes offer greater range of operations and at a time I to 4 or
sometimes 5 rows of holes depending upon the character of strata are drilled. The backmost
row of holes is usually along the toe of the high wall for the next succeeding cut and holes in this
row are frequently charged more heavily than holes in other rows, so as to facilitate digging of
the pilot trench. Quite often it is advantageous to adopt a differential spacing of holes, i.e., the
spacing between holes across the cut.
The dimensions of joints also affect hole spacing. The vertical holes terminate 0.60-
1.82 m. above the coal seam according to the nature of the seam. The efficiency of blasting with
vertical holes, however, will depend on the correct choice of the diameter of drill hole, the
burden, the spacing and on the choice of explosives and the manner of charging and distributing
the explosives in the hole.

2. Inclined Holes: As the inclination of the bench face is increased, the free face area becomes
more and hence with the increase in the inclination of the bench more explosive energy is
reflected back after the shock wave has reached the free face. As a result, greater energy will be
available for fragmentation of the rock. This results in better breakage of toe. Hence, for bench
blasting inclined holes are drilled in some situations.

▪ Advantages: The following are the advantages of inclined holes:

a. Better utilization of explosive energy and better fragmentation
b. Less back break
c. Easy toe-breaking
d. Less drill meterage required
e. High fragmentation results in cost reduction on account of loading, transport and
crushing.
f. More fragmentation and displacement are possible
g. Eliminate effectively front row toe burden and digging problem.
h. Probability of cut offs is lower
i. Benches of higher thickness can be worked effectively and efficiently
j. New face created by inclined hole drilling is smooth and sound
k. Produces less amount of overbreak
l. Produces less vibrations
m. Reduce secondary blasting cost 7 also reduce consumption of explosives
n. Cost of drilling, blasting, loading & crushing is reduced
o. Angle of inclined holes generally varies between 15-30 degrees

It is claimed that per 1° of inclination, 10% explosive energy is saved.

▪ Disadvantages: The disadvantage of inclined drilling are:

i. difficulty in drilling, and
ii. strain on the drill.
iii. Drilling accuracy decreases as the angle of drill holes increases and as a result there may
be the chances of much deviation in optimum burden and it leads to decrease in blasting
efficiency drastically
iv. Difficult in closer supervision. However, with the use of sophisticated drill machine, this
problem may be solved.
v. There is an increased risk of borehole collapsing
vi. Difficulty of loading explosives and booster charges into the blasthole.
vii. Difficult to put detonating fuse into the blasthole.
o Bench Blasting terminology: some of the major terms related to Blasting inn a bench, are as follows:
1. Bench Height: One of the primary factors that controls the design of a blast is that of bench geometry.
Usually, the bench height, H, is relatively constant for most multi-level pits and its value is set to conform
with the working specifications of loading equipment.
Bench height vary within wide limits. In large open pits from
which stone or minerals are mined, bench heights of 15-20m. are common, although benches with
heights up to 30m. are occasionally encountered. In many places, bench heights are limited as safety
precaution. In road building and other construction projects, the same work sites may have benches
varying from a few decimeters to several meters. The bench height is related to the degree of heaping
and spreading of material broken by blasting, thus, directly affecting displacement requirements to be
accomplished by the round design. The height also limits the maximum and minimum charge diameters
that should be used, and it influences drill selection. The maximum bench height is dictated by the
appropriate statutory authority. In general, faces with heights of about 10-18 m have been considered
the most economical and least hazardous to work. Where it is necessary to practice selective
mining/quarrying, the face height may be dictated by the thickness of ore/rock of a certain quality. The
most economical face height may also be determined by the drill penetration rate; when- ever the
penetration rate decreases significantly, it is generally uneconomical to drill deeper. High faces pose the
problem of considerable bit wander or drift, especially with smaller diameter blastholes. The deviation
of blasthole places a limit on the maximum allowable bench height.
In tests to determine limiting burden and spacing, without any consideration
of the degree of fragmentation, charge length (and hence bench height) was found to have a strong
influence on the limiting spacing between simultaneously initiated blastholes. The maximum spacing to
burden ratio for a given rock cannot be achieved unless a minimum charge length to burden value is
exceeded. When the charge is short (and the bench low), small changes in burden, B and/or spacing S,
have major effects. As the bench height and charge length increase, for a given burden, the limiting
spacing increases rapidly to begin with, then increases at a slower rate and finally attains a constant
value. The charge length should not be less than 3 times the burden if the maximum limiting spacing is
to be achieved with simultaneously-fired blast holes. In recent years, there has been a considerable
trend towards larger diameter blastholes. Because bench heights have either remained unchanged or
decreased slightly, a considerable increase in the stemming length to bench height ratio has been
brought about. At some open pit operation, for example, the use of large diameter blastholes (310 or
380 mm) and relatively shallow benches (12-14 m) prevent efficient charge distribution in the rock to be
fragmented; the rock alongside the stemming column (relatively remote from the charge) can be up to
40% of the total rock volume to be removed. From the view point of increasing blasting efficiency
(through reduced collar rock and greater charge length to burden ratios), there are good technical
reasons, where blasthole diameters are large, for increasing bench heights.

2. Bench Width: There is a minimum bench width, measured horizontally in a direction perpendicular to
the pit wall, for each bench height and set of pit operating conditions, whose value is established by the
working requirements of the loading and haulage equipment. The width also must be such as to ensure
stability of the excavation, both before and after blasting, because each blast effectively reduces the
restraint sustaining the pit wall at higher elevations. Because of the limits set by requirements for
equipment operating room and bank stability, there is a minimum bench width that should not be
exceeded by any blast.

3. Broken characteristics of materials: All materials expand when broken, and each has its own
characteristic swell factor, Sf. The factor is defined as the ratio of the volume of a unit weight of solid
material that when broken. The importance of swell in the design of blasts is that it directly affects the
choice of an initiation-timing system and limits the number of rows in the blast. For the box cut,
expansion will occur in two directions, while for the corner or side cut it will be in three directions. If
swell is assumed to be uniform, each solid dimension towards an open face could be expected to expand
to an amount equal in value to the original solid dimension divided by either the square root or cube
root of the S_{f} ratio characteristic of the material. The increase will vary from 5-30% with 15% being
 considered average. Loose material will not normally stand vertically of its own accord. Horizontal
spread towards all open ends beyond that due to swell will be a function of the natural angle of repose
of the material. Moisture and greater than normal throw from blasting will effectively lower the repose
angle and increase the spread.
It may be necessary to consider the effects of weathering on broken
material. Some materials deteriorate rapidly where exposed. In other instances, serious difficulties arise
during crushing, grinding, and screening of certain materials when they are wet. Therefore, total volume
and frequency of firing blasts will be limited by the maximum permissible exposure time of the material.

4. Operating restrictions: Each blasting pattern has distinctive effect on the heaping or scattering of broken
material, thus strongly, affecting loading. In the event shovels are used, the broken pile is best loaded
when its height is approximately equal to the maximum cutting height of shovel. Lower heights require
excessive shovel movement, while higher piles present the hazard of having to undercut boulders or
slabs that might be left on the pile top. If the ledge is lower than optimum, heaping is necessary to
provide good loading. In addition to height restrictions imposed by operating characteristics of the
loading equipment, there will be limitations as to spread of broken material over the bench floor. The
haulage method must also be considered in blast design. For example, conveyor and rail haulage
generally require long benches with material displacement during blasting being directed parallel to the
pit walls for a parallel loading approach.
Blasts necessarily should contain only a few rows of holes because
loading will be restricted quite often to only one side of the loader. Except for very high ledges, truck
haulage favour with several rows due to greater manoeuvrability of the haulage units. For draglines and
high lifts, greater displacement of broken material is desired than required for shovel.

o Blast Hole geometry: The major terms related to Blast Hole Geometry, have been defined under:
1. Blasthole Diameter: The hole diameter is selected such that in combination with appropriate positioning
of the holes, will give fragmentation suitable for the loading and transportation equipment and crusher
used. Additional factors to be considered in the determination of the hole diameter are: the size of
operation, the bench height, the type of explosive used and rock characteristics. Occasionally it may be
necessary to use two different hole sizes even within the same blast. As long as poor fragmentation does
not create problems for the mine operator, generally the larger the blasthole, the more economical
drilling and blasting become. The larger diameter holes could be drilled with only a small increase in the
cost per meter of drilling and larger holes could be put further apart. However, when the blasthole
diameter is increased and powder factor or energy remains constant, the larger blasthole pattern
generally gives coarser fragmentation. Perhaps by keeping burden unchanged and elongating spacing
alone the problem of coarser fragmentation can be overcome.
Where joints or pronounced bedding planes divide the
burden into large blocks, or where hard boulders lie in a matrix of softer strata acceptable fragmentation
is often achieved only when each block or boulder has a blasthole (Fig. 12.4). This usually necessitates
the use of smaller diameter blastholes and correspondingly smaller blast size. In a strata, which exhibit a
dense network of (natural) fissures, on the other hand, increases in blasthole diameter cause relatively
small change in fragmentation. The maximum possible charge diameter is rarely utilised since
fragmentation may be too large for efficient handling. The primary crusher and loader bucket sizes will
limit the largest size of material desired. The effects of rock structure sizing, how- ever, must be also
considered. Another factor that may restrict the maximum charge diameter may be when the loading
density of the explosive must be limited for ground vibrations or other adverse blast effects. In addition,
drilling time for the necessary meterage that satisfies production requirements may limit the size of
blast- holes for proper equipment operating balance.
The minimum charge diameter must be somewhat
greater than the critical diameter of the explosive because velocity of detonation increases as charge
diameter enlarges, but this is up to a certain finite value beyond which no further velocity in- crease
results. It is thus advisable to utilise that diameter where maximum velocity and energy yield per
 kilogram are attained. Hole diameters vary from 35 mm (13/8 in.) in small benches to up to 440 mm
(approx. 17 in.) in large benches. With small diameter holes the blast size is small. Large blasts need
large diameter deeper holes. In Europe and Scandinavia where rock is more competent, smaller
diameter holes 64-125 mm, are used on very high benches. In USA. Australia, Canada and India trend is
towards large diameter holes with lower bench heights. A combination of two conditions is practised in
different countries. Further, depending on type of rock different trends exist. For example, in India
limestone quarries use 100-150 mm diameter holes, in coal mines hole diameters vary from 150-269
mm, and in iron ore mines hole diameters used are generally 200-269 mm and above.

2. Blasthole Inclination: In recent years, increasing attention has been given by open pit and quarry
operators to the drilling of blasthole at angles up to 20° from the vertical. The benefits from inclined
charges are the reduction of collar and toe regions, less subdrilling requirement, and (usually) increased
throw. But air-blast and fly-rock may occur more easily due to the smaller volume of material
surrounding the collar. Most often it is advisable to reduce loading density in the collar zone to minimise
those effects. Inclined holes are successfully used in Europe and Scandinavia where high benches and
smaller diameter holes in medium to high strength rock are practised. The use of vertical blastholes
usually produces a considerable variation in burden between the top and bottom of the face (Fig. 12.5).
This is particularly the case where the face is high and/or highly inclined. Front row blastholes are often
collared near the crest so as to remove the heavy toe burden. But then, of course, explosion gases may
blow out prematurely in the upper face, causing high levels of noise, air blast and fly-rock. The rate at
which such venting reduces blasthole pressure near floor level may be sufficiently great to prevent
adequate breakage or movement of the toe. This effect is more pronounced for top-primed than the
bottom-primed charges. If a vertical blasthole is drilled at the nominal burden distance back from the
crest, on the other hand, hard immovable toe can be expected. One of the major advantages of inclined
blastholes, therefore, is the greater uniformity of burden throughout the length of the blasthole. Ideally,
the blasthole should parallel the face. For constant fragmentation, inclined drilling allows the use of
greater blasthole spacing. It also results in greater operating safety for men and machines due to the
cleaner face that is obtained. When blastholes are inclined, less subdrilling is necessary and so less
damage to the area beneath the pit floor is caused; drilling of the next bench therefore, becomes easier.
In some circumstances, there, may be very good reasons to go to the extra trouble of drilling inclined
blastholes.
Where benches are high, angles of 20-25° are recommended. Angles greater than
25° are seldom used because of difficulty in maintaining blasthole alignments, excessive bit wear and/or
difficulty in charging blastholes. If the blasting operation is to be successful, it is essential that the holes
are correctly drilled. To assure this, the holes must be properly aligned and straight. Deviation tends to
increase with increasing depth. Compensation for this deviation demands increased drilling. If the
amount of deviation is to be kept to minimum it is essential for the rod to be rigid. Full-section round
rods are more rigid than light rods with the same thread diameter. Deviation can be minimised by the
use of guide tubes, DTH drills using drill pipes and low feed force assure best hole straightness.

3. Burden: This is one of the most critical parameters in the design of blasts. The burden B, is the distance
from a charge axis to the nearest free face at the time charge detonates. With multiple row blasts, the
burden may not necessarily be given as the distance to the nearest free face. As boreholes with lower
delay periods detonate, they too create new free faces. As a result, the true or effective burden will
depend on the selection of the delay pattern. There are many relationships available for obtaining the
approximate value of the burden for various explosive and rock combinations. Most relationships utilise
either charge volume, charge weight or hole diameter as the basic parameter, with the bur- den being a
function of the cube-root or square-root of the independent variable. The cube-root law is stated as:
1
𝐵 = 𝐾 ∗ 𝑄3
where, B is the burden (ft or m)
 Q is the charge weight (lb or kg )

K is an empirical factor depending on the explosive and rock properties.

Of the many formulae using the charge diameter, De

as the independent variable, that proposed by Anderson (1952) was the simplest. The expression did
not consider either the properties of explosives or the rock. To account for the characteristics of the
explosives and the strengths of the materials, Pearse (1955) and later Allsman (1960) and Speath (1960),
proposed similar formulae. The difficulty in applying these relationships lies in the selection of
appropriate values for the properties of the explosive and rock. A modified formula was given by Ash
(1968) for the first ap- proximation for burden determination dependent on the borehole diameter:
𝐾 ∗ 𝐷𝑒⁄
𝐵𝑢𝑟𝑑𝑒𝑛 = 12
where B = Burden (ft),

De = Explosive diameter (in),

K = between 25 and 35, equal to 30 in average conditions, in rock with density around 2.7 g/cc.

If the rock density is higher than this, then K would be reduced and vice versa.

If the burden is too small, detonation gases escape into the

atmosphere in the form of noise and air-blast therefore less energy is available for the fragmentation.
Aside from the objectionable aspects, noise & air-blast are outward signs of the inefficient use of
explosives energy. Where the burden value is too large, gases are confined for a time interval longer
than desired; which can result in higher ground vibrations, excessive backbreak, toe and uneven floor. In
a ground, where good fragmentation is required and which is tough and/or blocky, the burden and also
the spacing should be conservative. When good fragmentation is less important, or when blasting
ground tends to break easily, satisfactory results may be obtained by drilling larger diameter blastholes
on a correspondingly larger pattern.

4. Spacing: The distance between adjacent blast holes, measured perpendicular to the burden is defined as
the spacing, which controls the mutual effect between holes. Spacing is calculated as a function of
burden, hole depth, relative primer location between adjacent charges and also depends upon initiation
time interval. The spacing is selected according to widely held concept that since the break angle made
by the charge to the bench face is near 90 degrees, hence spacing larger than two times the burden are
not possible (Gregory, 1973). Over the decades, inmost mining operations, spacing to burden ratios used
have been between one or two (Fraenkel 1957; Speath 1960; Kochanowski 1963) Commonly ratio
suggested is 1.2 to 1.3. Vutukri & Bhandari (1973) found from a survey of 100 operations that drilled
spacing to burden ratio practised was

𝑆 = 0.9𝐵 + 0.91
Where, both burden & spacing were in meters.

The works of Bhandari & Badal (1990) indicate that when the joints
are across the face, a close spacing is needed and when the joints are parallel to the free face larger
spacing ratios are possible. Whenever adjacent charges are initiated separately with a time-delay
interval of sufficient duration to permit each charge to break separately, there is no intersection
between the holes. In such cases, the spacing should be approximate to the Burden, i.e., B~S. When the
interval for initiating adjacent charges is reduced, there may be reinforcement of the stresses generated
by the Explosives in the zone between blastholes. Adequate results can normally be obtained when
spacing and burden are about equal forming a grid pattern. But the elongated patter, where spacing
exceeds the burden, is more effective, particularly in massive, hard-breaking formations. A large spacing
and small burden tend to cause more twisting and tearing of the rocks, less splitting along the line of the
blast holes, and less back break. Where short delay initiation methods are used, greater spacing are of
advantage in that there is less chance of cut-offs. Spacing appreciably less than the burden tend to
 cause premature splitting between the blast holes and early loosening of the stemming. Both these
effects encourage rapid release of gases to the atmosphere. In delay blasting of multirow shots, best
results with rectangular patterns are obtained at spacing two time the burden, and for staggered rows,
fragmentation improves as S increases up to about 4B.

Production increase up to 30% have been reported by decreasing B and

increasing S, values of blast hole length and charge weight/m3 rock remaining constant. With large S/B
values, the number of blast holes can be reduced and/or loading efficiency increased as a result of the
improved, more uniform fragmentation. Normally Spacing & Burden are related to Blast Hole depth and
more particularly to charge length. Changes inn Burden tend to affect the overall degree of
fragmentation and presence of toe or high bottom much more rapidly than changes in spacing. To
summarise, the following could be stated as basic principles for spacing:

1. For Sequence Delays in the same row, S~B

2. For simultaneous Timing in the same row, S~2B
3. For Multiple rows with sequence timing between charges in the same row, the entire round should
be drilled inn a square arrangement, particularly if identical timing is used for chares located laterally
with one another in adjacent rows
4. Staggered patterns are preferred between rows where all holes inn a single row are fired
simultaneously but timing between rows is delayed.

5. Sub Drilling: To avoid leaving a hump, bootleg or toe in bench blasting, the blastholes are drilled below
the floor (grade) level. This is termed as subgrade drilling, under-drilling or subdrilling. The optimum
effective subdrilling depends on:
a. The structural formation and density of the rock, -The type of explosive (and more particularly
the energy generated per meter of blasthole),
b. The blasthole diameter,
c. The blasthole inclination,
d. The effective burden, and - The location of initiators in the charge.

In practice subdrilling (J) should be roughly 0.3 times the burden, B.

If a pronounced discontinuity, or parting is located at grade level often no subdrilling is re- quired. Unless
such a condition exists subdrilling should be at least one-third the value of the burden. The required
subdrilling is also expressed in terms of the blasthole diameter, d. In dipping or massive rocks, subdrilling
of about 8d is usually found to be satisfactory. But where vertical blastholes are drilled in relatively high
and/or highly inclined faces, sub-drilling of 10d or even 12d may be necessary in front-row blastholes
because of the heavy toe burden. Subdrilling less than 8d can often be used satisfactorily when:

(1) A very high energy per meter of blasthole can be generated, and/or
(2) Bottom priming is done.

Too much subdrilling must be avoided, since it:

i) Wastes drilling and explosive expenditure,

ii) Increases ground vibration levels appreciably,
iii) Causes undesirable shattering of the pit floor (which may create drilling prob- lems on the
next bench down), and
iv) Increases the vertical movement of the blast.

In coal strip mines, subdrilling may in fact, have a negative

value. Where the overburden is shallow, blastholes are often stopped sufficiently short of the coal to
avoid shattering it in the blast. If soft beds lie immediately above the coal, good blasting results may be
obtained by drilling down only to the bottom of the lowest hard band in the overburden. But in most
cases, and especially where the overburden is deep, drilling is stopped as soon as coal appears in the drill
cuttings. If the coal is relatively weak, it may be necessary to backfill 1 m or so with drill cuttings to
 minimise shattering of coal by the overburden blast. Such shattering is particularly to be avoided where
coal seam is relatively thin.

6. Stemming: The primary function of stemming is to confine the gases produced by the explosive until
they have adequate time to fracture and move the ground. The type and length of stemming have no
significant effect on the characteristics of the explosion generated strain-waves and hence does not
increase stress-wave effect. By reducing premature venting of high-pressure explosion gases to
atmosphere, however, a stemming column of suitable length and consistency enhances fracture and
displacement by gas energy. Experiments have shown that the critical burden can be increased
significantly when a suitable type and quantity of stemming is used. For long charges, the effect of
stemming variations on the limiting burden is most significant for top primed charges.
The amount of unloaded collar required for
stemming is generally from one-half to two thirds of the burden dimension. This length of stemming
usually maintains sufficient control over the generation of objectionable air blast and fly-rock from the
collar zone. When the burden has a high frequency of natural cracks and planes of weakness, relatively
long stemming columns can be used. When the rock is hard and massive, the stemming should be
shortest which prevents excessive noise, air-blast and/or backbreak. The resistance to ejection of water,
mud and wet clay, stemming depends almost entirely on the inertia of the column. Dry granular
materials, on the other hand, exhibit both inertial and high frictional effects. For this reason, dry granular
stemming is much more efficient than materials which behave plastically or which tend to flow.
Where rock
is hard and massive, the stemming column should be the shortest which prevents excessive noise, air-
blast, fly-rock and/or backbreak. The use of small 'pocket' charges may be considered. In multirow
blasts, where the mean direction of rock movement tends more and more towards the vertical with
successive rows a longer stemming column is often used in the last row to reduce overbreak (Fig. 12.14).
In some mining fields for small diameter shallow hole blasting, the availability of cheap ANFO blasting
agents has led to the complete filling of a hole with ANFO and no stemming is used. In this case, it is
presumed that gases have self-stemming effects.

7. Powder Factor/Specific Charge: Two terms are often used to relate explosive mass and consequent rock
broken: Powder factor and specific charge, q. Observation of blast designs based on empirical relations
has been often indicated in terms of powder factors. Powder factor (sometime also referred as charge
factor) is the ratio between the total weight of ex- plosive detonated in a blast divided by the amount of
rock that is broken. It is usually expressed as kilogram per ton or kilogram per meter. In some cases,
reciprocal of these are referred as powder factor but correctly these are termed as specific charge.
These are sometimes termed as charge ratio also. As the powder factor in kg/m³ increases, the average
fragment size decreases when the burden, B, remains constant. According to Gustafsson (1981), when B
exceeds 3 m, the fragment size becomes uncontrollable, especially in the upper portions of the
blasthole, where charge densities are the lowest. The powder factor varies between 0.1 kg/m³ and 0.53
kg/m³ for bench blasting. Powder factor for tunnelling should be larger by 1.25-1.5 times then bench
blasting because of the larger fixity.

S. Major Removal Equipment Predominant Powder Factor Bench Height

No. Geological Unit (kg/m3) (m)
Fragmented
1 Dragline Shale 0.30 15.0
Shale 0.35 23.0
Shale 0.40 -
Sandstone 0.35 18.0
2 Small Dragline Shale 0.20 6.0
Shale 0.50 7.6
Shale 0.35 18.0
 Sandstone 0.65 26.0
3 Front End Loader Shale 0.35 9.0
Sandstone 0.65 16.0
Sandstone 0.40 20.0
Sandstone 0..95 15.0

The following Table 12.1 compares various powder factors for

surface mines when vibration is not a consideration. Table 12.1 also relates powder factor to muck
removal equipment and geology for surface coal mining with ANFO. The smaller the muck removal
equipment, the larger the powder factor necessary to obtain desired fragmentation.

In order to obtain good fragmentation and thereby

ease in loading operation, the explosive consumption in excavation is somewhat greater than in
quarrying. When firing is confined to single row of blastholes in soft laminated strata, the charging ratios
may be as low as 0.15-0.25 kg/m³. In harder sedimentary strata the charging ratios generally are around
0.45 kg/m³ while they may be about 0.6 kg/m³ in jointed igneous rock. Even higher charging ratios may
be necessary in order to obtain satisfactory results in some metamorphic rocks such as mica schist,
which absorb much of the energy of the blast. Generally, 1 kg of explosive will bring down about 8-12
tons of rock.

8. Energy Factor: Till the introduction of slurries, water-gels and emulsions; the powder factor was a good
indicator of the amount of energy used to break a quantity of rock. However, with slurries, and
emulsions, the energy can vary greatly when small charges are used in the strength tests and also when
density remains constant. There- fore, many use a different method to relate the amount of explosive
energy required to fragment a given quantity of rock. This is termed as the Energy Factor. The amount of
theoretical energy is used as an index of effectiveness of an explosive to break rock. Many different
computer programmes are available which determine the theoretical energy yield. Thermochemical
energy is expressed in terms of calories per unit volume or weight of an explosive. The Absolute Bulk
Strength (ABS) of a given explosive is the amount of thermochemical heat energy expressed in terms of
calories per cubic centimetre. The Absolute Weight Strength (AWS) of a given explosive is the amount of
thermochemical heat energy expressed in terms of calories per gram. The Energy Factor (EF) can be
defined as the amount of explosive energy kilo calories and its distribution relationship in a given
quantity of rock. Therefore,
𝑘𝑖𝑙𝑜 𝑐𝑎𝑙𝑜𝑟𝑖𝑒𝑠
𝐸𝑛𝑒𝑟𝑔𝑦 𝐹𝑎𝑐𝑡𝑜𝑟 =
𝑞𝑢𝑎𝑛𝑡𝑖𝑡𝑦 𝑜𝑓 𝑟𝑜𝑐𝑘𝑠
The energy factor can be related to either
cubic meters or tons. There is a difficulty. However, that different computer programs using exactly the
same input and output characteristics, give significantly different strengths on the same composition.
Thus, one has to be careful while using these theoretical calculations given by different manufacturers.

o Factors to be considered while blast designing: are as follows:

1) Bench Height (HH): Normally dictated by site parameters which include type and dimensions of loading
equipment. If the height is not predetermined, then it should be greater than or equal to 2.5*Burden
Distance. While drilling bench heights greater than 4*B, care is to be taken due to hole deviation. The
typical values for large scale operations is 10-15m. and that for small scale operations, it is kept between
6-12m.
2) Hole Diameter (D): It depends on drilling equipment. For optimum fragmentation use, hole diameter
equal to the bench height divided by 120. The maximum hole diameter should be equal to the bench
height divided by 60. When using charge diameters that are less than the hole diameter, the effect of
decoupling must be taken into account. Smaller holes distribute the explosive energy better than larger
holes. Cost of drilling, loading and explosives per hole volume generally decrease wit increase inn hole
diameter. For Large scale operation, it is kept between 250-350mm. and that for small scale operations,
it is kept between 83-200mm.

3) Explosive Type: are as follows:

a) ANFO: Holes must be dry or lined. Low Cost. Simple to mix, safe to handle components. Safe in use.
Low bulk strength. Good for low powder factor situations. Maybe expensive in High powder factor
situations owing to high drilling costs.
b) Slurry/Emulsions: Used in dry or wet holes. Usually mixed and pumped by contractors. Higher Bulk
Strength and hence lower drilling cost than ANFO, but lower weight strength & higher overall cost
than ANFO.
c) Heavy ANFO: Advantages of both ANFO & emulsions. Cost advantages and can be used in wet or dry
conditions. Possibility of varying strength characteristics.
d) Cartridge Explosives: Whenever the quality of ANFO is poor or site mixed products are not available,
these are used. Labour Intensive. Costly Products. Cartridges are needed for controlled blasts.

4) Burden (B): Based on the characteristics of rocks & explosives to be used, burden is determined by using
appropriate relationship or experimentally determined from Bench Crater Method.
𝐵 = 0.024𝐷 + 0.85 𝑚.
It is generally kept from 20-35 *
explosive diameter (D). Existing Rock Hardness, Fractures, Explosive used and the required
fragmentation dicates the burden.
a) Minimum allowable distance from the crest to front row: Determined by Safety requirements, Drill
Dimensions and Bench Crest conditions
b) Front Row Mean Burden: Crest to Front Row Distance + ½ Bench Height * Cotangent Face Angle
c) Front Row Mean Burden Volume: Mean Front Row Burden * Spacing * Height
Generally, for Large Scale
operations, its value is kept between 5-10m. and that for small scale operations, it is kept between 3-7m.

5) Burden / Spacing Ratio:

a) Square Pattern: Easy to Layout, Poor Charge Distribution, More difficult to Delay. Generally, 1-2
b) Staggered Pattern: Even Charge Distribution, easy delaying, Good in large patterns for low powder
factors. Generally, 1.1-1.4
c) Wide Spacing: Good for high powder factors in competent rocks Not suitable for joint orientation
across the face. Delay Sequence will alter drilled burden and spacing. Recommended Effective Ratio.
Generally, 1 to 3-4

6) Spacing determination from Burden: Generally, for Large Scale operations, its value is kept between 6-
11m. and that for small scale operations, it is kept between 4-7m.
𝑆 = 0.9𝐵 + 0.91 𝑚.

7) Stemming (T): Inert material placed in the hole in collar to confine gasses. Crushed rock confines
explosive energy better than drill cuttings. Usually equal to 0.7 to 1.5 time the burden dimension. If the
stemming is less than ~ B, then Fly-Rock and premature venting may occur. Generally, for Large Scale
operations, its value is kept between 2.8-7m. and that for small scale operations, it is kept between 2-
5m.
8) Subdrilling (J): The distance drilled below grade level is calculated by the formula,
𝐽 = 0.3 𝑡𝑜 0.5 × 𝐵𝑢𝑟𝑑𝑒𝑛
If 0.5 * Burden still leaves an excessive toe,
then the burden distance should be reduced. No subdrilling is used if hole through the overburden ends
at the top of coal seam. Generally, for Large Scale operations, its value is kept between 1-3m. and that
for small scale operations, it is kept between 1-2m.

9) Hole depth (L): Check the hole depth in relation to the rock burden and bench height.
𝐿 = 𝐻+𝐽
Generally, for Large Scale operations, its value is
kept between 12-17m. and that for small scale operations, it is kept between 8-12m.

10) Broken Volume/Hole: Generally, its value is kept between 500-2000m3. It is calculated form the
following formula:
𝐵𝑢𝑟𝑑𝑒𝑛 × 𝑆𝑝𝑎𝑐𝑖𝑛𝑔 × 𝐵𝑒𝑛𝑐ℎ 𝐻𝑒𝑖𝑔ℎ𝑡

11) Hole Inclination: Blast Hole should be in conformity with the face. Vertical to 20 degrees uniform burden
and need less subdrilling.

12) Estimation of Powder Factor: Controlled by Rock Conditions. Maybe determined by empirical
relationship. Design powder factor should be expressed in terms of weight or volume of standard
explosive It is generally kept between 0.25-1.2kg/m3 or 0.1-0.4kg/t

13) Actual Powder Factor: Measured for whole blast or total production-less than design powder factor due
to overbreak and free digging.

14) Explosive Charge/Hole (Q): Generally, it is kept between 350-700kg. and is calculated by the following
formula:

Density of explosive * Available hole volume (Cross sectional area of hole) * (Bench Height + Subdrill -
stemming)

15) Charge Distribution:

a) Bottom Charge is the amount of explosive required to fragment the toe region. It is generally kept
between 20-40%.
b) Column Charge is the amount of explosive required to fragment the remaining rock associated with
the hole. It is generally kept between 80-60%.

16) Drill Pattern: Select a pattern of drill hole that is estimated to be able to provide the desired rock
movement. Take into account any possible environmental problems that may limit the maximum
permissible charge weight or the number of holes fired per delay interval. Mark the sequence of charge
initiation on the plan, select an appropriate number of rows and the number of holes per row.

17) Delay Interval: 3 to 6 ms/m of effective Burden. Place delays 1m. from delayed holes. The typical values
of it are 17, 25, 35 or 45 ms.

18) Delay Sequence: delay to Free Face. Delays must be used to avoid choked blast. Delays must be used to
modify spacing: burden ratios. Delays can be used to reduce maximum instantaneous charge level.
 19) Controlled Blasting: Apply Necessary controlled Blasting Techniques. Back Row to final wall distance
should be equal to the anticipated back break. Diggable back-break approximately equal to burden.
Varies with rock conditions, explosive strength and powder factor of back row.

20) Number of Holes per Charge: Depends on the quality of rock to be broken, length of face and spacing,

21) Number of Rows: Depends on the delay time, type of initiation system and the rock structure.

o Single & Multiple Row Blasting: In an opencast coal mine, both the vertical and inclined holes parallel with
the bench face is practised. Rows of the hole may be in single or multiple. Firing Patterns in Opencast Mines,
may be classified into two main groups:
▪ Single Row Blasting: In Single Row Blasting, the choice between firing the holes simultaneously or in
a sequence depends on the nature of the rock and the degree of fragmentation and throw desired.
Delay firing of the series of holes in a row is done to obtain better fragmentation. The other
advantages of delay firing are the reduction in ground vibrations, less back-break or over-break, and
better control of the rock pile. The delay interval is obtained by using short delay detonators, or
detonating relays in conjunction with detonating fuse. Two basic types of short delay blasting
patterns for one row are illustrated in Figure A & B, and for better fragmentation, the pattern shown
in figure C may be used.
Multiple - row blasting pattern gives better fragmentation but
specific explosive consumption is lower & hence more economic. Because of this reason, in most of
the coal benches and coal measure rocks, Single Row Blasting is preferred. Single Row with short
delay firing is generally implemented in small mines with narrow working berms of benches and also
to reduce the overcrushing of the products. However, This type of firing generally yields more
oversized boulders and less amount of yield compared to Multi row Blasting with short delays. In
single row firing, the following patterns may be used:
▪ The alternate delay pattern: It is used for soft rock formation
▪ Consecutive Short Delay Pattern: It is used for rock with medium hardness
▪ Short Delay Firing with a cut: It is applied for hard rock formation. The cuts are made by
drilling a number of cut holes at a specified distance from the face. The cut holes are over-
drilled by 1 to 1.5m as compared to the other holes. It has the following advantages:
▪ It will provide additional free faces
▪ It assists to break the rock boulders by collision among themselves
▪ Assists to form the muck pile in particular direction
▪ It ensures good fragmentation
The demerits associated with it are:
▪ More amount of back break
▪ Difficult to work in toe
▪ More ground vibrations

▪ Multi Row Blasting: In multi row blasts, while in-line firing systems are also used, a large number of
other patterns are available which provide better fragmentation and blasting efficiency and varying
the effective burden BE and Spacing SF. In initiating multi-row rounds of holes, two factors are
important, which are:
▪ The point where the breakage begins on the first row as shown by a in figure D should as far
as practicable be located where either the rock structure is weakest or the burden s most
favourable, and
▪ The inherent ignition scatter or detonators of the same delay period should be taken into
account while designing the ignition pattern so that every hole has a free breakage zone in
front as illustrated in fig. E.
 In multi row firing, the two basic short delay firing
for single row, can be applied in various ways. A few of ignition patterns have been illustrated. The
pattern shown in Fig. D gives better fragmentation. A modification of this pattern is shown in Fig. F,
which gives better scope for free breakage of the apex holes. All the holes in a row except the edge
holes can be ignited with one and the same delay where by any hole that happens to be ignited first
will have perfectly free breakage whichever it may be. In the case of benches with unlimited width,
the ignition pattern can be simpler and more reliable, even when one interval per row is employed.
It is possible to avoid the construction at the end holes by making the blasting finish diagonally to
the face as shown in Figure G.
▪ Transverse Cut Pattern: They are used where smaller width of muck pile is desired
▪ Wedge or Trapezoidal Blasting Pattern: These are used when the rocks are medium hard
and hard one. Due to the motion in the opposite directions in this case, the big boulders are
broken by supplementary collision. This is better than the parallel or V cut since the trunk
line does not form acute angles, offers good fragmentation, reduces percentage of
formation of boulder less than 0.5%, does not form toe, achieves better throw, reduces
secondary blasting and since firing burden is less than the actual burden, it offers better
fragmentation.
▪ Diagonal Blasting Pattern: With this type of blasting pattern, it is possible to blast the rock
towards the least resistance and improve the fragmentation of the rock.
▪ Parallel Cut: In case of bench length is very high but bench width is limited for only 3 rows of
blastholes, then this pattern of cut is being followed.
▪ Flat Face delay Pattern: It is indicated in Figure. The number against each hole shows the
order of detonation by the ms delay detonator and the arrows show the direction of
movement of muck piles. This type of delay pattern is not very much effective for achieving
the good fragmentation. This type of pattern can be used with the square, rectangular or
staggered drill pattern.
▪ The ‘V’ Pattern: This type of cut is used if the bench length to width ratio is 3:2. In this cut,
try to avoid acute angles to prevent misfires/cut-offs. Number of Rows should be limited to
3 or 7 and the total firing time is limited to 3000ms.These types of patterns, are useful for
any type of rocks and may be used with square, rectangular or staggered hole pattern.
However rectangular or square grids are more common. In case of rectangular grid spacing is
more compared to burden. The various types of ‘V’ pattern are shown in Figure.
▪ The Echelon Pattern: Whenever a 2nd free face is available in the outside corner and when
better fragmentation and displacement are required then this pattern is introduced. It has
been shown in Figure.
The Multi Row Short Delay Firing improves the quality of
fragmentation, lowers down the number oof holes to be blasted, makes easy the working in a large
reserve block, proves the opportunity for concentration of work and makes easier for scheduling of
complete drilling, blasting and loading operation.

▪ Choice: Decision of choice among the Single Row or Multiple Row Drilling and Blasi=ting depends
mainly upon:
▪ Firing Sequence
▪ Quality of Fragmentation:
▪ Shape of the muck Pile wanted:
▪ Factors of working on Bench Toe

o Transport of Explosives in Bulk: This practised is used for Transport of explosives for Deep Hole Blasting.
▪ Rules: The rules related to it are as follows:
▪ The Owner, Agent or/and Manager of the mine where transportation of explosives in balk is
proposed to be practiced, shall ensure the following.
▪ Only properly trained persons who are authorised in writing by the manager for the purpose
are deployed.
▪ Adequate personal protective equipment as required to be used by persons deployed in this
connection are provided and also used.
▪ The entire operations of transportation of explosives in bulk within an opencast mine shall
be placed under the overall charge of a competent person holding at least an Overman's
certificate of competency.
▪ Transport of explosives in bulk to the priming station or the site of blasting shall be done
only during day light hours.
▪ The quantity of explosives to be transported in bulk at one time to the site of blasting shall
out exceed the actual quantity required for use in one round of shots, and also not before 30
minutes of the commencement of charging of holes.
▪ Only a vehicle duly approved by the Competent Authority shall only be used for transport of
explosives in bulk.
▪ All conditions stipulated by the licensing authority in respect of the vehicle deployed for
transportation and handling of explosives in bulk shall be strictly followed
▪ Such vehicle shall be in safe operating condition and should be driven by competent licensed
drivers duly authorised by the Manager. At least two fire extinguishers of suitable size and
capable of fighting electrical and petroleum fines shall be provided in each vehicle in an easy
accessible position and maintained in a state of readiness.
▪ Before transporting explosives in bulk, the competent person authorised in this regard shall
personally search every person engaged in the transport and use of explosives and shall
satisfy himself that no person so engaged has in his possession any cigar, cigarette, bidi, or
other smoking material or any match or any other apparatus like mobile phone etc., of any
kind capable of producing a light, flame or spark.
▪ Additionally, the following precautions shall be strictly observed while transporting
explosives in bulk
▪ The vehicle shall be properly earthed with chain links while loading
▪ The vehicle shall be well locked except during times of placement and removal of
stocks of
▪ The vehicle shall not be overloaded.
▪ The vehicle shall not be driven at a speed exceeding 25 kilo-meters per hour.
▪ The vehicle loaded with explosives shall not be left unattended.
▪ The vehicle shall be kept in isolated places while loaded.
▪ The vehicle loaded with explosives shall not be taken into garage or repair shop and
shall not be parked in a congested place.
▪ The vehicle transporting explosives shall not be refuelled except in emergencies;
even then it's engine shall be stopped and other precautions taken to prevent
accidents.
▪ Wherever, drilling operations are being carried out, charging of already drilled deep
holes shall not be carried out in the same area at the same time.
▪ Every vehicle used for the transport of explosives in bulk shall be carefully inspected once in
every 24 hours by a competent person, to ensure that:
▪ fire extinguishers are filled and are in place,
▪ the electric wiring is well insulated and firmly secured.
▪ the chassis, engine and body are clean and free from surplus oil and grease,
▪ the fuel tank and feed lines are not leaking and
▪ lights, brakes and steering mechanism are in good working order.
▪ A report of every inspection made under sub-clause (a) shall be recorded in a bound paged
book kept for the purpose and shall be signed and dated by the competent person making
the inspection.
▪ The mine manager shall frame a suitable code of practice for handling and transportation of
explosives in bulk.
 ▪ Precautions: are as follows:
▪ Transport of explosives from the magazine to the priming station or the site of blasting shall
not be done except in the original wooden or card board packing cases. The quantity of
explosive transported shall not exceed the actual quantity required for use in one round of
shots. The explosive shall be transported to the site of blasting not more than 90 minutes
before charging the holes.
▪ No mechanically propelled vehicle shall be used for explosive transport unless that is
approved by the Chief Inspector. Every vehicle used for the transport of explosives shall be
marked, on both sides and ends with the word "EXPLOSIVES and shall be provided with at
least two fire extinguishers (one of carbon tetrachloride type for petroleum fire and the
other of carbon dioxide type for electrical fire). The vehicle used for explosive transport shall
not be overloaded.
▪ Jeep or land rover may be used for transport of detonators from magazine to priming
stations subject to the following conditions
▪ not more than 200 detonators are transported at a time
▪ the detonators are packed in wooden box;
▪ the wooden detonator box is placed inside a metal case approved by the Chief
Inspector;
▪ the metal case shall be suitably bolted to the floor of the vehicle;
▪ no person shall ride on the rear portion of the vehicle.
▪ Explosives and detonators shall not be transported in the same vehicle at the same time.
▪ No person other than the driver and his helper (not below 18 years of age) shall ride on a
mechanically propelled vehicle. The vehicle loaded with explosive shall not be left
unattended.
▪ Before unloading or left standing the engine of a vehicle shall be stopped and the brakes
shall be set securely. The speed of vehicle shall not exceed 25 km./bout. The vehicle shall
not be taken into garage or repair shop and shall not be parked in a congested place.
▪ The vehicle transporting explosive shall not be refused except in emergencies when engine
shall be stopped. No trailer shall be attached to a vehicle carrying explosive.
▪ The vehicle used for the transport of explosives shall be inspected once in every 24 hours by
a competent person to ensure that:
▪ fire extinguishers are filled and in place;
▪ the electric wiring is well insulated and firmly secured,
▪ the chassis, engine and body are clean and free from surplus oil and grease;
▪ the fuel tank and feed lines are not leaking, and
▪ lights, breaks and steering mechanism are in good working order.

A report of every such

inspection shall be maintained.

▪ All operations connected with the transport of explosives shall be conducted under the
personal supervision of a foreman solely placed in charge of blasting operations at the mine.
▪ The blaster shall search every man engaged in the transport and use of explosives and shall
satisfy that no person so engaged has in his possession any cigarette, bidi or other smoking
apparatus or any match or any other apparatus, capable of producing a light, flame or spark.

o Secondary Blasting: While performing deep hole blasting, sometimes big boulders are left behind, which
create a problem for the loading machines and crushers. The secondary blasting is done to break over-sized
boulder, produced during primary blast, to size, suitable for handling and the crushing plant. There are two
basic methods of secondary blasting:
▪ Pop Shooting: For Pop Shooting, the boulders are to be selected first and shall be marked for drilling
holes. A hole Is normally drilled just beyond the centre of the boulder to be broken. The depth of the
 hole shall be around 2/3rd or 3/4th of the size oof the rock. With some kinds of rock shot holes 30 cm
deep are sufficient to break very large boulders.
The number of holes to be drilled depends upon the size,
shape, type and placement of rock boulders. The charge varies with the size of the boulder. The cap-
sensitive cartridge explosives are inserted in it (generally 200 gm of explosive for every cubic meter
of the boulder) and fire after taking due precautionary measures. The shots are fired either by a
safety fuse and plain detonator or by electric detonators. It is necessary to withdraw mechanical
loading equipment to a safe distance, since there is considerable scatter of rock when firing Pop
Shots.
The charge varies with the size of the boulder but, for average conditions, a
boulder 90 cm * 60 cm * 60 cm requires a charge of about 130g of Special gelatine-80% strength. If
electric shot-firing is practised, the electric detonators are connected in series and the shots are
fired simultaneously. The following are the major specifications:
▪ Number of holes per boulder 1 or 2 depending on Boulder size
▪ Depth of Hole 2/3rd of boulder thickness; Min. Depth-0.5m.
▪ Charge per Hole 60-100gm Special Gelatine
▪ Dia. Of hole 35mm.
▪ Initiation Device Ordinary Electric Detonator
▪ Muffling Procedure Sand Bag covering each hole
▪ Expected Volume of Rock/Blast Hole 2m3

▪ Plaster shooting: The plaster shooting provides a ready means of breaking even large boulders in
circumstance where drilling is difficult or expensive. A high-velocity, high-strength gelatine type
explosive is most suitable for this work. A charge of Plaster Gelatine primed with a detonator and
safety fuse, or an electric detonator is laid on the surface of the boulder, it is then covered with a
shovelful of plastic clay which is pressed into position by hand. It is advantageous to wet the surface
of the stone before plastering, and the clay should be well pressed down so that it is in good contact
with the surface of stone around the explosive.
In Plaster Shooting, the charges used are about for times those
required for Pop Shooting. The advantages of this method are as follows:
▪ No drilling is required, thus saving labour and compressed air, a consideration of particular
importance in hard and abrasive rocks which may be difficult to drill. For the same reason,
even though the amount of explosive used is much greater than in pop shooting, it is found
to be more economical in overall cost.
▪ A group of plaster shots can be prepared far more quickly than the same number of pop
shots.
▪ Boulders are broken where they stand and the fragmented material is not scattered over a
wide area.
▪ There is less likelihood of damage through flying debris. This is advantageous since
 equipment need not to be moved far to a place of safety and thus less loading time is
wasted when secondary shots are fired.

o Controlled Blasting: When the area near the blasting zone (within 500m.), consists of surface structures,
railway lines, industrial buildings, etc. the issues like Ground Vibrations and Fly-rocks, etc. are removed by
the proper selection of blasting parameters. This is known as Controlled Blasting. Increasing use of
techniques to control damage to the remaining rock is being practised in surface mines.
▪ Objectives: The objectives of controlled blasting are as follows:
▪ To prevent over & under fragmentation
▪ To reduce Ground vibrations
▪ To reduce noise
▪ To reduce dilution or waste of ore
▪ Reduce fractures within remaining rock walls
▪ To confine the blasting area
▪ To reduce the fly-rocks.
▪ Types: The major techniques included in this category are as follows:
1) Presplitting: Presplitting creates an artificial discontinuity along the periphery of the designed and
expected limit of the main blast fragmentation to isolate the blasted zone from the remaining rock-
mass. It helps to prevent overbreak and creates smooth and stable highwall, and also help reflection
of shock waves from the subsequent main blast.
This is primarily used in quarrying. Some of the shock waves also get
reflected, thus hindering the ground vibrations henceforth. Pre-splitting can be done in two ways:
a) by in hole delay arrangement and
b) by air deck method of presplitting holes.

• Air Decking Technique: When an air-gap is introduced in the midst of an explosive columns,
explosion products from two portions of the charge generate shock waves (Fig. 44, 45, 46)
which move in opposite directions. In the middle of the air-gap there will be a collision of the
shock waves. As a result of this collision, a new source of high pressure is formed in the
centre of the air- gap. After reflection, the shock waves will change direction and start
moving towards the bottom of the hole and the stemming. Reflected from these hard
obstacles the shock waves will again collide in the centre of the air-gap and the process will
be repeated. This results in better utilisation of explosives energy and uniform
fragmentation throughout the length of the borehole.

In this technique, the air gap is introduced by using

special air bags, that contains chemical materials in it. When the air bag is rubbed with hand,
the chemical inside, it starts reacting. It is then immediately lowered into the hole at its
desired position, with the help of a rope. Due to the chemical reaction, the bag inflates by
the gases created in the reaction, and gets fixed into the desired position. This inflatable
plug seals a borehole at any point to separate the stemming from an air gap and explosive
toe load. The air decking technique using inflatable power plug can be used for ensuring pit
wall stability and selective mining of coal seams by predrilling under seam waste wedges.
This shall be the better technique in case of badly fractured ground. The crack formed by the
presplitting should be at least 5 mm in width. General procedure in air deck method of
presplitting is as following:

▪ First estimate the burden and spacing by measuring the tensile strength of the in-
situ rock-mass.
 ▪ Drill holes of 200 to 250 mm diameter up to a depth varying from 10 m to 30 m as
per the design requirement. Spacing in between the holes may be 5 to 12 times the
diameter of the holes. The distance of the production hole is generally 3.5 times the
diameter of the hole.
▪ Pour ANFO with primer charge up to a depth of 1/10th of the depth of hole.
Explosive charge generally 0.9 to 1.3 kg/m² of split.
▪ Cover the holes with the power plug near to the mouth of the holes.
▪ Stem the holes say up to a height of 3 m with the stemming material as mentioned
earlier. Length of the stemming column should be at least 3.5 to 4.0 times the
diameter of the hole.
▪ Blast nearly 50 m length of the bench at a time.

As an alternative, a wooden spacer can be used in place of

the air bag. The air bag technique is being used in Jambad OCP, Parez OCP, Chasnala OCP,
Nigahi OC, Murali Hills Limestone mines and some other manganese mines. The wooden
spacer techniques are being used in some non-coal quarry.

• Multideck In-hole Delay Technique: In multideck in-hole delay blasting the arrangement is
done in the following manner:
▪ First estimate the burden and spacing of holes by measuring the tensile strength of
the in-situ rock-mass.
▪ Drill holes of required diameter up to the design depth.
▪ Insert primer explosive first into the hole along with primer
▪ Detonating Relay Cords (DRC) may be used for firing if only one delay is to be
provided for blasting a hole. DRC may be available at 15, 20, 25, 42, 50, 100, 150 ms
delay periods
▪ After pouring first primer explosive, stem the holes with stemming material.
▪ Insert primer explosive 2nd time over the stemming material along with primer.
▪ Stem the remaining portion of the hole with stemming material.
▪ Continue to charge the holes by alternate stemming and primer explosives as per
the design.
▪ Blast the holes.

Multideck in-hole delay blasting will reduce the

vibration level and maximise charge per delay (Fig. 48).

• Advantages of Presplitting: are as follows;

▪ The presplitting will eliminate the back breaks on the high wall side,
▪ there will be increase in permissible charge 300 to more than 400 times
▪ reduction in ground vibration level more than 70%.
▪ Presplitting also stabilize the highwall,
▪ ensure high percentage of coal or mineral recovery,
▪ ensure safety to both man and machine working in the mine
▪ extends safety to inhabitants in the close proximity of surface coal mines.

2) Deck Charging: When the overburden of opencast is composed of alternate hard & soft bands, the
explosive charge is located opposite the hard bands instead of placing the major portion of the
charge at the base of the hole. The intervening space between charges in a hole are filled with
stemming (consisting in most cases of drill-hole gumming) and the top of the hole is stemmed up to
the Collar. This method of charging is known as Deck Charging. The deck Charging is used to reduce
the kg/delay of explosives, so that the ground vibrations are less.
 This method will yield less oversized lumps and also it will have smaller zone of rock breaking.
Generally, in the lower part of the hole, 2/3rd of the charge of high strength and density explosive is
placed and the remaining 1/3rd charge may be divided into one, two or three supplementary charges
in the middle and in the upper parts of the holes. The explosive for the supplementary charges may
be lower strength and low-density type. The number of decks and the length of the holes depends
upon the minimum permissible length of stemming (which may be reduced by 15-25% in this case).
The Length of the air gap between the deck may be equal to the 20 to 35% of the length of the
charge or distance between the deck charge is filled with stemming material of stemming length 60
to 80% of the length of the charge. The smaller length of the air gap is used for blasting stronger
rocks. Here initiation is done not with the charge, but by an intermediate detonator with the help of
a detonating cord.

• Advantages: are as follows:

▪ The material is
neither blasted too
fine, nor it forms
boulders.
▪ The ground vibrations
are less.

3) Chamber Blasting: Generally, this type is blasting is practised in the very high benches of soft rock
formation, to reduce the total meterage of drill holes spaced at sufficiently large distances. Since the
maximum charge is required at the toe of the benches, the size at the bottom of the hole is enlarged
by firing one or two small chamber charges. Then, the total block is blasted like normal blasting
method after placing the charges in all the holes made inn the ore block. This type of blasting may
also be practised in the very hard rock formation, where holes are made by Jet-Piercing and bottom
of the hole is reamed with special type of drill bits.

4) Coyote Blasting: In this system a heavy amount of explosive charge in the order of hundPlasterreds
of tonnes are packed in the large one or more coyote chambers made by driving tunnels, drifts or
raises, to excavate a huge amount of rock in the order of thousands of tonnes from a hilly terrain or
rock body where sufficient out of free faces are available for the movement of dislodged rock. After
packing explosive in the Coyote Chambers, the connection tunnels/drifts/raises near the chambers
are backfilled tightly with the excavated muck-piles and the charge is normally blasted with the help
of detonating cord.

5) Cushion Blasting: It is also called the Trim Blasting. In this system a heavy amount of closely spaced
and burdened (same dia. as the production hole) vertical holes are drilled all along the periphery
(beyond the last rows of holes in the bench) of the blasting zone in the bench. They are loaded with
light explosive charge and nicely stemmed and are blasted after the blasting of the main charge with
the help of delay detonator. Deck charging may also be practised in this type of blasting. Generally,
the diameter of the explosive cartridges ishalf the diameter of the blasthole (2-4 inches). The
cartridges are generally lowered into the hole by attaching with the detonating cord at
predetermined intervals. The remaining air space in the hole is completely back filled with the
stemming material.

The blast holes are fired after the rest of the round has been blasted
Generally, nitro-glycerine-based cartridges explosives are used for this controlled blasting technique.
The diameter of this cartridges is less than 50% of the diameter of the blasthole. The cartridges are
generally spaced by a string although in some of the cases they may form a continuous column.
Higher density explosive is put at the bottom of the hole. The diameter of the hole may vary from
37.5mm to 150mm, the diameter of the cartridge varies from 22mm to 50mm (depending upon the
size, spacing and burden of the blast holes), spacing of charges varies from 100mm to 550mm,
approximate weight of charge varies from 0.14 Kg/m to 1.47 Kg/m length of the blasthole, spacing of
boreholes varies from 0.6 to 2.1m, burden of boreholes varies from 0.9 to 2.7m, etc. All the holes
meant for cushion blasting are fired simultaneously using a detonating fuse trunk line. In some of
the unconsolidated rock formations upper half of the holes are lightly charged to prevent back
breaking. This method is simple and economical, as the number of holes drilled are less (thus, the
spacing between them is more).
6) Line Drilling: Heavily Charged holes (away from the periphery) which provides the primary blasting is
termed as the primary holes or production holes. Uncharged holes are drilled along the periphery of
the blasting zone. Spacing between these holes is kept around 2-4 times the hole diameter. These
uncharged holes cause s some of the shock waves to reflect. Lightly charged holes adjacent to the
uncharged holes are called buffer holes (50% less explosive than primary blast holes and 50-75%
closer as compared to them). This technique gives the maximum protection to the host rock. It is
costly, as the number of holes to be drilled is too high.

7) Smooth Blasting: It is also called Contour Blasting. In this method, the contour holes are drilled
along the next excavation line and are lightly loaded than the buffer and production holes. Spacing is
kept loser than buffer and production holes. The contour holes are fired simultaneously to produce
excavation contour without fracturing the adjacent rocks. Generally, 10-12 times hole diameter in
soft rocks, are kept as spacing.

8) Muffle Blasting: This method is mainly employed to tackle Fly-rocks. This is based on the principle of
covering the blasting zone. A wire mesh or a blasting mat is laid on the blasting zone. Steel wire
expandable nets can also be used, to spread continually over the holes having the same delay
number. Gunny Bags, filled with sand or drill cuttings, weighing 0-59 kg are kept over the mesh/mat
at an interval of 3-4m. The empty containers of explosives, each weighing around 30 to 50 kgs. Can
also be used as an alternative. Wire mesh is generally made of steel ropes.

The wire mesh prevents the fly-rocks from

flying, like an umbrella. In earlier days, this technique was used for blasting within 100m. from any
surface structures. However, it was only useful for less dia. And less depth holes, and not for
production blasting. With the advent of Shock-tube and electric detonators, the fly-rocks can be
 prevented with the help of proper buren-spacing. The Muffle blasting is pnly used for road
construction, metro rail construction etc. nowadays.

• Precautions: are as follows:

▪ Used tyres are kept below the wire mesh, to prevent the wires of it from breaking.
▪ The wire mesh is kept carefully, without scratching or breaking the wires of
detonators.

9) Deep Hole Blasting: The blasting done in opencast mines, in which, the depth of blast holes is
greater than 3m., is called Deep Hole Blasting.
• Precaution while charging & firing of holes: are as follows:
▪ Before conducting the drilling, the area shall be checked to be free from any socket
or explosive.
▪ The drilled holes shall be sealed with the help of a plug, before being charged with
explosives.
▪ The Red Flags shall be erected in the blasting area.
▪ Only one hole shall be charged at a time. There shall be no person on the bench, at a
radius of 20m. from the hole being charged.
▪ The charging, stemming and connection etc. shall be done, by standing at a solid
place, away from the face.
▪ Warning sirens shall be fired, in order to warn the person of the blasting. Guards
shall be placed at conspicuous places.
▪ The firing shall be done, when the Overman makes sure that all the machines have
been retreated to safe places.
▪ The task of priming, charging and stemming shall be done under the supervision of
the Overman. He shall himself conduct the blasting, from the shelter.
▪ All the mobile phones shall be switched-off while performing the tasks of Explosive
handling, charging or firing.
▪ If during the task of charging or firing, lightning occurs, then:
✓ Explosives/detonators shall not be handled, and shall be not left at an open
place.
✓ If hole charging is being done, the work shall be stopped.
✓ If charging has been done, the end of detonating cord shall be covered with
a tape, and the mouth of the holes shall be covered with a wood or stone.
Workers shall be retreated to safe places.
✓ If connection has been done, the workers shall be immediately retreated to
safe places.
 • Precaution while drilling and blasting of deep holes: are as follows:
▪ The are where holes are to be made, shall be made clear of any loose boulders. The
overman shall mark the position of every hole, on the ground, with the help of tape.
▪ The explosives shall be kept at a place, that is far away from the place of priming,
and shall be at such a place, that people do not have to travel near to them.
▪ The blasting shall be done in the daylight and shall be done on the same day, on
which, the holes have been charged.
▪ As far as possible, the task of blasting shall be done during the interval of shifts, or in
between the two shifts.
▪ Before 30 minutes from the blasting, the danger zone shall be marked with the help
of a red flag.
▪ Before 10 minutes of firing of holes, the warning sirens shall be made in the pattern
that has been predetermined by the manager of the mines.
▪ If the danger zone has any railway line, or roadway, no blasting shall be done, before
stopping the traffic.
▪ After shot-firing, the Blasting Overman shall inspect the blasting face, after an
interval of at least 5 minutes (or until the smoke and dust settles). If he finds
everything OK, he shall make the sirens for “All Clear” and then only the normal
working be resumed.
▪ The sirens shall be such, that they can be clearly heard at a distance of at least
500m. in all directions. All the workers and the people living nearby shall have a
clear information regarding the Danger & All-Clear warning signals.
• DGMS instructions regarding Controlled Blasting: These have been mentioned in Circular
General 2 of 2003:
▪ The are within 500m. from the place of blasting shall be considered as a Danger
Zone. If it is required for a person to be present in the anger zone, during the period
of blasting, he shall take a Blasting Shelter.
▪ The manager of the mine shall form a Cide of Conduct for Controlled blasting, that
shall consist of precautions to be taken while performing the blasting through:
✓ Milli-second delay detonator, non-electric shock tube, sequential blasting
machine
✓ Muffle blasting
▪ If the Danger zone consists of residential and industrial buildings, it shall be ensured
that the ground vibrations does not exceed, 5 & 10 mm/sec. respectively.
▪ A proper training shall be given to the person performing controlled blasting and the
person supporting them.

Extras:

1. Haul Roads: These are also called the life-line of Opencast Mines. The reason behind this is that the cost of
transportation of Overburden and Coal, in an opencast mine, is the greatest as compared to other task in the
mine. Any problem in the design and maintenance of the Haul Road, will directly affect the productivity of
the mine, and affects its working. In this way, the haul roads are considered the lifeline of opencast mines.
The haul roads in an opencast mine shall be made by keeping various parameters in the mine. Moreover,
they shall also be maintained properly and frequently. This will increase the production of the mine, and
shall lower the costs related to the maintenance of the machineries.
a. Effects of Poor Haul Road: The following are the demerits of a poorly maintained haul road:
i. The life of the dumper reduces
ii. The cost of diesel increases to a considerable extent.
iii. The machineries need more frequent maintenance.
iv. There is a constant possibility of big accidents.
b. Parameters of Good Haul Road: The following points shall be borne in mind while maintaining the
haul roads:
i. Width: The width of the Haul Road shall be at least three time the width of the widest
machinery plying on it.
ii. Gradient: The gradient of the haul road shall not exceed 1 inn 16. The gradient of the ramp
can be increased up to 1 in 10.
iii. Curve: the following points are considered:
1. Horizontal Curve shall be neglected as far as possible.
2. Horizontal and vertical curve shall not be made at a same site.
3. The gradient of the curve shall be kept uniform.
iv. Sight Distance: At a curved road, the driver shall be able to see clearly up to a distance of at
least 30m.
v. Super Elevation: At every Horizontal Curve, a superelevation shall be provided according to
the speed of the dumper.
vi. Drainage: A drainage of proper size and gradient shall be provided on both sides of the Haul
road.
vii. Parapet Wall: Wherever there is a possibility oof a dumper falling from the side of the haul
road, a parapet wall of 1m. height shall be provided on the side of the Haul Road.
c. Maintenance: The maintenance of a 5km. long Haul road requires:
i. Equipment: following:
1. Dozer – 1
2. Grader – 1
3. Back-Hoe – 1
4. Road Roller – 1
5. Water Tanker – 1
6. Compacter – 1
ii. Manpower: following:
1. Civil Engineer – 1
2. Civil Supervisor – 2
3. Dozer Operator – 1
4. Dozer Helper – 1
5. Back-Hoe Operator – 1
6. Back-Hoe Helper – 1
7. Road Roller Operator – 1
8. Road Roller Helper – 2
9. Compacter Operator – 3
10. Water Tanker Operator – 1
11. General Mazdoor – 10
12. Total - 24