WALL CONTROL BLASTING TECHNIQUES

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					                   WALL CONTROL BLASTING TECHNIQUES
   TO MINIMIZE DAMAGE TO THE ROCK AT THE LIMITS OF SURFACE AND
  UNDERGROUND EXCAVATION, IN ORDER TO ENHANCE SAFETY STANDARD
                         AND ECONOMY

                                                     ***

        Author: Partha Das Sharma, B.Tech(Hons.) in Mining Engineering,

                                 E.mail: sharmapd1@gmail.com,

              Blog/Website: http://miningandblasting.wordpress.com/
                                                   ABSTRACT

 Wall failures are costly and often life threatening. The goal of efficient wall control blasting is to make
transition from a well fragmented rock mass to an undamaged slope in as short a distance as possible. This can
be quite challenging due to the many factors that influence wall damage. To develop efficient designs one must
have a basic understanding of wall failure mechanisms as well as limitations of wall control procedures. In
addition, it is imperative, design be precisely implemented, evaluated and refined on a continuous basis. The
release of energy during blasting produces reactive forces, which cause the deterioration of the remaining rock
face. Pre-splitting and trim blasting are the key techniques adopted to protect final rock faces. However, even
these well known techniques are applied; slope failures and back damage may persist. The key parameters
within the control of the blasting engineers are type and energy in the hole, drilling pattern, hole depth, hole
diameter, hole angle, bench geometry and blast timing. An understanding of mechanisms of all the aspects is
needed for good designing for blast for wall control and slope stability.


1. INTRODUCTION:

Wall control blasting is the technique used to obtain a pit wall, free of backbreak and loose rock that
will stand safely at the required wall angle for extended periods of time. Direct damage to the
excavation limit due to blasting is usually found in the form of backbreak or overbreak, crest fracture
and loose rock on the face. The mine operator has a number of tools available for minimizing or
eliminating these problems. Techniques include changing the explosive type, or changing the
blasthole diameter, by decoupling the explosive, by decking, and by changing the burden and
spacing. Changing the depth of subgrade drilling or the stemming height can reduce crest fracture
and any resultant narrowing of the width of safety benches. Changing the millisecond delay timing
and the rotation of the round may also be helpful in eliminating these problems.

The rock characteristics and geology must be considered when designing controlled blasts as these
have an important influence on the final results. The compressive strength, crushing strength and
tensile strength of the rock should be known. The frequency and orientation of joints and fractures
in the rock are also important parameters. These variables cannot be controlled but must be
determined by suitable field and laboratory techniques.




                Author: Partha Das Sharma, (B.Tech-Hons., Mining Engg.),
E.mail: sharmapd1@gmail.com, Website: http://miningandblasting.wordpress.com/                           Page 1
WALL CONTROL BLASTING TECHNIQUES
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In other words, wall control blasting techniques have been used in surface and underground blasting
in the mining, quarrying and construction industries for many years. The specific reasons for the use
of controlled blasting techniques may vary according to the industry and project; however, two
generally applicable reasons can be identified:

a. To insure that the rock is broken to the excavation limit but not beyond to keep the host rock
intact.

b. To insure the subsequent safety of personnel and equipment, working under / side of the wall, by
avoiding back-break and loose rock on the face.

In open pit operations breakage beyond the pit limit is costly. Excessive back-break at the perimeter
generally results in an overall pit wall angle less than designed, and may result in the need for costly
artificial support techniques. In fact, failure to properly control blasting at the final pit wall can cost a
large open pit mine many millions of dollars expenses in additional waste removal for the same ore
mined.

There are four principal controlled blasting techniques used, which are:

• Pre-splitting

• Cushion blasting

• Buffer blasting

• Line drilling

Pre-splitting is the most commonly used technique especially in surface work. This is followed by cushion
blasting, also known as trim blasting in open pits. Smooth blasting, used underground, is similar to
cushion blasting.

Pre-splitting provides a preferential fracture plane behind the blast to terminate cracks growing from
blast holes, while trim blasting reduces rate of energy release against the final wall.

Buffer blasting may be used alone in cases where the rock is quite competent, but this is not a common
approach. However, a properly designed buffer row at the back of the final production shot is essential to
the success of most pre-splitting and cushion blasting applications.

Line drilling involves the drilling of closely spaced small diameter holes at the perimeter of the excavation.
These holes are not loaded with explosive, but form a discontinuity at the excavation limit. This method is
costly because of the many boreholes drilled and is therefore only seen in blasting for civil works projects,
where back-break can be a very expensive result. Modified forms of line drilling may be used in mining
and quarrying in special circumstances.

Geology can have pronounced effects on the results of controlled blasting / wall control blasts. For
example, it is known that trim blasting does not work well in the presence of relatively shallow dipping
joint planes dipping into the excavation. It may not always be possible to obtain the classic result when
adverse geology is encountered. However, if back-break, crest fracture and face loose rock have been
minimized, then the result will be far more acceptable than a wall in the same rock where no controlled
blasting has been performed.


                Author: Partha Das Sharma, (B.Tech-Hons., Mining Engg.),
E.mail: sharmapd1@gmail.com, Website: http://miningandblasting.wordpress.com/                          Page 2
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Underground, over-break in the stope results in costly ore dilution. Poor breakage control at the
perimeter of drifts and shafts means more scaling of the walls and roof and more difficulty installing
support and facilities.

In construction blasting breakage beyond the designed limits may lead to the removal of many tons
of rock not specified in the contract. Added scaling and support may be needed for the long term
stability of the wall. The consumption of concrete and other construction items may well increase.
All of this is expensive.

Equally important as cost, in every industry, is the need to provide a safe working environment. Pit
and quarry walls that have sustained substantial back-break are prone to hazardous rock falls. Safety
benches, intended to arrest the fall of loose material will typically be narrow and ineffective. Drifts
and stopes experiencing excessive over-break will be more prone to hazardous rock falls. Similar
hazards will also exist in construction work as well. Therefore, any organization that emphasizes
safety will want to control blasting at the limits of an excavation.

Thus, wall control blasting techniques are the system of controlled blasting which refers to various
techniques used to minimize damage to the rock at the limits of an excavation due to the action of
the ground shock wave and the high pressure explosion gases, generated during the blast.

2. GENERAL PRINCIPLES OF WALL CONTROL BLASTING TECHNIQUES:

a. Controlling energy input given by explosive and the borehole pressures exerted - A fundamental
goal of all wall control blasting is to reduce the energy input and the borehole pressures at the
perimeter of the excavation. The borehole pressures generated by commercial explosives, which are
fully coupled to the hole, are much greater than the rock strength and will cause extensive damage
around the blasthole. Therefore, these pressures must be reduced.

The borehole pressure for a fully coupled hole can often be obtained from the manufacturer of the
product being considered for use.

CALCULATION OF BOREHOLE PRESSURE:
Borehole pressure can also be calculated using the following formula given. Generally, borehole
pressure is function of VOD of explosives used.
(Pb)c = 2.5 x 10-6 x ρ x V2 ;
where, ‘(Pb)c’ is borehole pressure in kilobar, when fully coupled explosive used,
‘ρ’ is density of explosives and
‘V’ is Velocity of Detonation (VOD) of explosives in m/s.

While the above equation may not yield exact results it has proven quite adequate for practical
design requirements. However, the equation has some limitation in the case of aluminized explosives.
The velocity of detonation is reduced because the initial reactions of the oxidizer with aluminium are
endothermic. However, beyond the detonation zone the equilibrium shifts to the very rapid formation
of exothermic reaction products. Therefore, the actual borehole pressure will be considerably higher
than that calculated from the detonation velocity.

Low density explosives produce low borehole pressures because the detonation velocity is reduced.



                Author: Partha Das Sharma, (B.Tech-Hons., Mining Engg.),
E.mail: sharmapd1@gmail.com, Website: http://miningandblasting.wordpress.com/                   Page 3
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Decoupling and decking are the techniques basically used for reduction of borehole pressure. A
primary means of reducing the borehole pressure is to decouple the charge from the hole. This
means that the diameter of the explosive charge is less than the diameter of the hole. Pressure may
be further reduced by decking, whereby gasbags, wooden or cardboard spacers etc., are used
between charges or the charges are taped to detonating cord with a gap left between individual
cartridges.

For a given hole diameter and explosive the usual approach is to decouple radially first. If this is
insufficient to reduce the borehole pressure enough, additionally decks can be employed.

 When a charge is decoupled from the blasthole the explosion gases expand to fill the hole volume
before exerting borehole pressure. Therefore the decoupled borehole pressure will be much less
than the coupled value.



CALCULATION OF COUPLING RATIO AND DECOUPLED PRESSURE:

Coupling Ratio (CR) can be expressed by:

CR = (C)1/2 x ( dc/dh),

where ‘C’ = the percent of explosive column actually loaded,

‘dc’ = charge diameter and

‘dh’ = hole diameter.

Decoupled pressure may be calculated from the following formula:

(Pb)dc = (Pb)c x (CR)2.4

where ‘(Pb)dc’ = The borehole pressure for a decoupled and/or decked charge,

‘(Pb )c’ = The borehole pressure for a fully coupled charge and

‘CR’ = Coupling ratio.

In using these equations it is necessary to have an idea of what an acceptable decoupled borehole
pressure will be. In pre-splitting it has been found that the pressure should be in the range of 2 to 5
times the uniaxial compressive strength. In larger hole diameters it is often better to set the
decoupled borehole pressure near to the uniaxial compressive strength of the rock because of the
greater radius of rupture that may result around larger diameter boreholes, when the borehole
pressure exceeds the compressive strength of the rock. This potential for large rupture radius around
the borehole can lead to a wall more prone to unravel over time.



Discussion on Borehole pressure on type of controlled blasting - In some presplitting applications a
concentrated charge is used in or near the bottom of the hole with the remainder of the borehole
left void. Upon detonation the explosion gases are free to expand up the hole and exert a suitable
decoupled pressure on the surrounding rock. This method has been used extensively in active
highwall pre-splitting when blast casting in dragline mines. It has also been used in other types of

                Author: Partha Das Sharma, (B.Tech-Hons., Mining Engg.),
E.mail: sharmapd1@gmail.com, Website: http://miningandblasting.wordpress.com/                     Page 4
WALL CONTROL BLASTING TECHNIQUES
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mining, generally being most successful if the ground is reasonably competent thereby avoiding
damage at the bottom of the hole and excessive leakage of gases as these expand up the borehole.

b. The buffer row - Occasionally buffer blasting alone may be sufficient to protect a final excavation
limit from damage. However, when pre-splitting or cushion blasting the last row of the final
production blast must be a buffer row. The exceptions to this rule would be when active highwall
pre-splitting for a dragline operation or in small diameter work underground where a buffer row is
not always used.

The buffer row must be designed with a sufficient charge to break the rock between the buffer hole
and the final wall. However, the explosive consumption in the buffer row should not be so great as
to cause breakage beyond the plane of the final wall or the controlled blasting effort would have
been wasted. Often, when damage is observed beyond the final wall limit the problem is the buffer
row design rather than the presplit or trim row.

As buffer row is designed with less explosive in the hole than is in production blasting boreholes;
because the explosive is kept low, in the hole, with a greater length of stemming above, there is less
potential for crest fracture and face loose rock. The toe between the buffer hole and the excavation
limit can still be adequately broken.

The low centre of gravity of the charge in the buffer hole causes it to behave like a spherical charge,
for which cube root scaling applies. In a buffer row a scaled depth of burial (SDOB) of about 1.5
times the optimum scaled depth of burial for the given explosive in the given rock type should be
used. The scaled depth of burial is simply the depth from the surface to the centre of the charge
column divided by the cube root of the total explosive weight in the column. Ideally the charge
should have a length not exceeding 8 times the diameter of the borehole. If, because of the hole
depth or diameter, the charge length exceeds 8 times the diameter the calculation should be
performed using the depth to the centre of a charge column equal in length to 8 times the diameter
and located at the top of the charge.

c. Effect of water on a decoupled explosive charge - When a decoupled charge is surrounded by
water the pressure generated by the detonating explosive, at the borehole wall, will be considerably
higher than would be the case if the explosion gases were free to expand across an air filled gap. The
degree of decoupling achieved will be much less than that calculated assuming the charge is
surrounded by air. In fact because water is quite incompressible the pressure transferred to the
borehole wall may be quite similar to that of a fully coupled hole. The explosive charge will need to
be decoupled to a greater extent than normal. If the area can be dewatered prior to final wall
blasting this will be the best solution.

When height of water column is above the concentrated presplit charge at the bottom of a large
diameter hole, another problem can develop. The water column tends to behave as stemming and
the explosion gases are inhibited from freely expanding up the hole. There will be more damage
around the bottom of the hole. The presplit crack may not extend the full length of the borehole.
These holes will work best if pumped before explosive loading. They should be loaded and fired
promptly to minimize the water column that forms above the explosive charge.



                Author: Partha Das Sharma, (B.Tech-Hons., Mining Engg.),
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3. INFLUENCE OF SITE AND STRATA CONDITIONS:

The properties of the rock and the site geology are of significant importance when designing a
controlled blast. If these factors are ignored the results will be, at best, a hit and miss affair. Serious
backbreak, crest fracture, face loose rock or sliding of weak portions of the wall are all possible
outcomes.

It is also important to recognize that in complex geological settings it may not be possible to achieve
the classic result. However, even though the half-casts of all the holes are not visible on the face the
controlled blast will still have been successful if a safe, stable wall has been achieved at an
economical cost.

a. Important rock properties to be considered - The most important rock properties are the tensile
strength, compressive strength and crushing strength. Also very important are the nature, frequency
and orientation of joints and fractures, the rock density, longitudinal wave velocity and Young's
Modulus. Ideally these properties should be measured in-situ. In-situ values reflect the effects of
weathering and structural features in the rock. A rock which tests as quite strong in the laboratory
may be considerably weaker when weathering, groundwater alteration, presence of structures such
as open joints, bedding or foliation planes and fractures due to previous blasting are accounted for.
However, at this time methods for measuring rock properties in-situ are not very satisfactory and
are usually costly.

Therefore, laboratory tests are generally relied on. Laboratory data can be adjusted by a site factor
to account for in-situ conditions. Deciding what the site factor should be is not a simple task and will
be an approximation.

Most practical is to design the controlled blast based on the laboratory results and observe the
results in the field. Then the design can be adjusted to account for any problems until an optimum
result is obtained. It may then be possible to back calculate the in-situ uniaxial compressive strength
and tensile strength.

Backbreak and radial crushing around the borehole result when the stress produced in the rock by
the explosion exceeds the crushing strength of the rock. The crushing strength is typically two to five
times the uniaxial compressive strength. Major backbreak problems are likely if an explosive loading
that was successful in competent ground is subsequently used in highly jointed or fractured ground,
even though the rock type is the same. Therefore, powder factors and decoupled borehole pressures
must be adjusted to account for structural conditions and the actual crushing strength of the rock
surrounding the hole.

Another point one has to observe is the joints orientation. The orientation of the joints has a major
influence on the controlled blast results. When joints or fractures strike parallel to the excavation
face a smooth clear wall may be obtained. When the joints are steeply dipping (>70°) the wall can be
made to conform to the joint planes.

When the joints are shallow dipping it is undesirable to cause the wall angle to conform to these
planes. There is greater chance that planes will undercut the face. When this occurs it is more
difficult to obtain a classic result because there is a greater likelihood that portions of the wall will

                Author: Partha Das Sharma, (B.Tech-Hons., Mining Engg.),
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WALL CONTROL BLASTING TECHNIQUES
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slide off along these structured planes. Large diameter cushion blasting has been found unsuited to
these conditions. Presplitting may be more successful if great care is taken to design the presplit and
buffer rows to minimize the disruption experienced on the joint planes. It takes relatively little
movement along the plane to destroy cohesion resistance and cause the material resting on the
joint to be more prone to slide.

When steeply dipping joints dip back into the wall while striking parallel to the face, sliding on
undercut planes is not possible. However, toppling failures may occur. In the presence of these
features the final wall should not be vertical. An angle of 70 to 80 degrees is more suitable. A toe
buttressing effect is provided and the wall is far more likely to remain safe and in good condition for
the long term.

When structural features strike at angles other than parallel to the face the amount of backbreak
depends on the nature of the joints and fractures and their strike. Open joints are likely to break
back more than tight, infilled joints. Planes striking at 45 degrees to the face are likely to break back
further than near vertical joints striking at 90°.

Again, the frequency of jointing is important. Jointing begins to interfere with wall control results
when the joint spacing is less than the hole spacing. In pre-splitting the hole spacing should not
exceed twice the major joint spacing. Frequent jointing can lead to greater crest fracture. The
explosive collar height must be increased or the upper column load reduced.

When the stress due to the reflected ground shock wave at the free face, near to a blast, exceeds
the rock tensile strength slabbing can occur. If joints, bedding planes or foliations exist, striking
parallel to the face, the potential for slabbing is greatly increased. Slabbing is especially a hazard
when blasting near to tunnels or when blasting in a pit that is in close proximity to the walls of
another pit. Reduced explosive loading may be necessary for better result.

Where rock breakage is not desired, as in the case at the final excavation limit, rock properties that
relate to the in-situ rock strength are important. The Young's Modulus of Elasticity is a measure of
the brittleness of a rock and its susceptibility to backbreak. Rock with a high Young's Modulus has a
higher crushing strength and is harder to break. Higher borehole pressures may be permissible at
the perimeter.

Rocks with a higher longitudinal wave velocity are also usually found to be stronger. Weaker rock or
strata that have been weakened by weathering, alteration or fracturing due to dense jointing or
previous blasting exhibits a lower longitudinal wave velocity. This fact leads to the seismic
techniques for determining overburden depth, depth of broken rock, radius of rupture, jointing and
density. As an in-situ method these techniques may be particularly valuable for determining the
nature of the in-place rock.

4. WALL CONTROL PRACTICES IN SURFACE OPERATIONS:

a. Explanation of various methods –

(i) Buffer Blasting - This is perhaps, the simplest form of wall control shooting. The last row of the
production blasting pattern is altered to limit the energy input at the final wall. The explosive loading

                Author: Partha Das Sharma, (B.Tech-Hons., Mining Engg.),
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WALL CONTROL BLASTING TECHNIQUES
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is reduced and as a consequence the burden and spacing are also decreased. As described above,
explosive loading is often reduced by selecting a scaled depth of burial greater than would normally
be used. Another approach is to use decoupled bagged powder above a toe load of fully coupled
explosive.




Buffer blasting can only be used as the sole controlled blasting technique when the ground is quite
competent. Some minor backbreak or crest fracture may develop but this will be much less than
would be caused by the production blast holes. Where buffer blasting can be used alone the cost of
wall control will be quite economical.

In most cases buffer blasting is used in conjunction with another wall control blasting technique. A
properly designed buffer row is very important to most successful presplit or trim blasts. Design of
the buffer row is the same as when the technique is used alone. It becomes important to insure that
the buffer row is at the correct location relative to the presplit or trim row.

Typical design for the buffer row includes using a scaled depth of burial at the top of the charge of
1.5 times the production hole value and reducing the powder factor to 0.5 - 0.8 times the production
row powder factor. Burdens range from 0.5 to 0.75 times the production burden. The spacing should
not be less than the burden and will usually be 1.0 to 1.25 times the buffer row burden.


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To avoid backbreak and crest fracture the buffer row holes must be properly located in front of the
intended plane of the final wall or the presplit line. This distance must be sufficiently large to insure
that the stress at the final wall is adequately attenuated to avoid crushing beyond the plane of the
wall. Following figure shows as an example, how the stress generated by detonating buffer row
holes attenuates with distance from the blasthole.




 It has been observed that, in quite soft rock, such as coal mine overburden, spacing the buffer row
10 feet (about 3.5 m) or more in front of the presplit line may indeed be prudent. In hard rock the
spacing at the toe needs to be much less to break the rock between the buffer row and the presplit
line. However, breakage beyond the presplit can be avoided.

Moreover, to avoid crest fracture in competent rock, drilling the presplit holes on an angle is
advantageous. One can space the presplit and buffer hole closely at the toe for breakage while
obtaining a greater standoff at the crest.

In hard rock it has been found that the toe of the buffer row should be 3 to 5 feet (1 to 1.5 m) from
an intended face angled at 80 degrees. In soft rock, such as coal overburden, it has been necessary
to move the toe of the buffer row out as much as 15 feet (4.6 m) to keep the zone of crushed
material from extending beyond the planned wall location.




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(ii) Presplitting - Presplitting is the most common controlled blasting technique and has proven
successful in applications from large open pit mines to civil construction. This method involves the
drilling of closely spaced holes at the planned excavation perimeter which are lightly loaded with
explosives in order to generate an appropriate borehole pressure. Presplitting is being done using
hole diameters ranging from 2 inches to 12¼ inches. Often, small diameter presplitting is preferred
for technical reasons and because the cost per square foot of wall may be lowers. Other mines use
large diameter holes in order to employ the same drills for presplitting as for production drilling. This
approach has worked especially well in active highwall pre-splitting designs associated with blast
casting operations. It has not always been as successful in other types of mining applications.

In small diameters (<5 inches, 127 mm) spacings of 3 to 6 feet (0.9 to 1.8 m) have been common.
When the decoupled borehole pressure can be permitted to significantly exceed the rock
compressive strength, then spacings of 7 to 9 feet (2.1 to 2.75 m) have been used successfully in 3-
inch (76 mm) boreholes, greatly reducing the cost of wall control. In larger diameter (>6 inches, 152
mm) hole spacing of 5 to 18 feet (1.5 to 5.5 m) have been employed. As spacing become larger
geological structure becomes an increasingly important control on this dimension.

     •    Spacing Between Holes - The spacing between the holes is a function of the hole diameter,
          decoupled borehole pressure and the tensile strength of the rock. The spacing between
          presplit holes may have to be varied in different areas of the pit if differing rock types exist
          with different uniaxial compressive strengths and tensile strengths. Therefore
          characterization of the geology is important. Not only do the rock properties affect the
          spacing, but the geological structure is also an important control. As a rule of thumb the hole
          spacing should not exceed twice the spacing between major, open joints.
     •    Pre-splitting on an Angle - Observations in open pit mines and quarries has shown that pre-
          splitting at an angle less than vertical contributes to a wall that remains in better condition
          for extended periods of time than one that is presplit vertically. This has been observed in
          iron mines, coal mines and quarries. Vertical presplit may be appropriate where the rock is
          particularly competent, or special circumstances preclude an angled wall. Presplit angles
          typically range between 70 and 80 degrees, with 80 degrees being perhaps the most
          common. In construction blasting a vertical presplit is likely to be more common. An angled
          wall may lead to greater construction cost. However, in deep road cuts for example an
          angled presplit should still be considered. A principal advantage to angle hole pre-splitting
          results from the toe of the presplit face being moved out from the crest. Therefore, if
          isolated blocks of rock fall from the face near the toe the entire face is not undercut, as
          would typically be the case for a vertically presplit wall. Another primary advantage occurs
          when steeply dipping joints or bedding planes dip back into the wall and strike near parallel
          to the face. Under these conditions the wall may be subject to toppling failures. The stability
          of a wall prone to these failures can be enhanced by the toe buttressing effect of an angled
          presplit wall. The third important advantage to angled presplit holes relates to the relative
          position of the presplit and buffer rows. When the presplit holes are angled and the buffer
          row is vertical it is possible to locate the toe of the buffer hole close to the presplit line for
          good breakage, while maintaining a greater standoff at the crest to avoid excessive crest
          fracture.


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WALL CONTROL BLASTING TECHNIQUES
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     •    Choosing the Hole Diameter - Current open pit and quarry designs call for multiple benches
          to be brought back to the final limit between safety benches. In general it is not possible to
          drill an angled hole flush to the wall using large hole equipment. Small diameter percussive
          drills, however, can perform this task quite readily by drilling back under the machine.
          Therefore, these machines are commonly used where the above criteria are to be met. In
          some cases a larger diameter drill may be used to produce the angle presplit, as in blast
          casting operations for example, if there is sufficient clearance room for the drill to set up on
          the holes. Again, the use of small diameter holes is not appropriate if the boreholes are
          quite deep. The limit is about 50 feet (15.2 m) on hole depth, although 60 feet (18.3 m) is
          possible in highly competent rock. In heavily fractured ground 40 feet (12.2 m) is likely to be
          the maximum depth to which small diameter holes can be accurately drilled. Also, in very
          wet ground small diameter holes are more difficult to drill with the desired degree of
          accuracy, if the holes are more than 40 feet (12.2 m) deep. Increasing the decoupled
          borehole pressures beyond the compressive strength of the rock has been more successful
          in small diameter holes than in large. The radius of rupture around a smaller hole is less.
          Therefore, any cracking that occurs around the borehole is less likely to cause long term
          unraveling of the wall of the excavation. From the spacing equation one can see that an
          increased decoupled borehole pressure results in a wider spacing between presplit holes,

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WALL CONTROL BLASTING TECHNIQUES
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          thereby reducing the cost. Thus the cost of small diameter presplitting will not always
          exceed the cost incurred using large diameters as is sometimes believed. Each situation
          should be assessed according to the factors discussed above and the best option selected.
     •    Shooting the Presplit Line - The presplit line may be shot with the final production blast or
          before the final shot is laid out in the field. Both approaches are workable. When the presplit
          line is detonated with the final blast it should be initiated approximately 100 milliseconds
          before the final wall blast. In delayed blasts care should be taken that the presplit line does
          not precede the detonation of the adjacent buffer row holes by too great a time. A delay
          may need to be introduced into the presplit line periodically in order to avoid the possible
          disruption of nearby buffer holes from the detonating presplit holes. However, as many
          holes as possible should be shot instantaneously taking into account the lead time and any
          vibration control requirements, because this yields a better defined presplit. If the final wall
          shot is quite narrow the presplit row should be detonated with the final blast. Detonating
          the presplit holes in advance may lead to the mass of rock sliding off the wall, leaving very
          poorly fragmented material to be cleaned up. Following figure is an example of a final wall
          blast incorporating two production rows, a buffer row and the presplit holes angled at 80
          degrees. This example is for an iron ore mine in competent rock.




     •    Active Highwall Pre-splitting in Dragline Operations - The pre-splitting technique has also
          been used to control the successive highwalls in a blast casting operation. The standard
          method involves drilling large diameter holes on the designed highwall location and loading
          these with a concentrated charge of explosive in or near the bottom of the borehole. Active
          highwall pre-splitting has two advantages. First, it allows a very regular highwall to be
          produced. Therefore, front row burdens on the next casting shot can be well controlled for

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          maximum casting efficiency. Second, in wet ground the presplit, fired in advance of drilling
          off the production blast, can be used to dewater the block to be shot thereby reducing
          explosives cost. When dewatering is a goal the presplit row will be drilled along the back and
          both sides of the block to be shot to isolate the area from recharge by groundwater. Active
          wall pre-splitting has often been accomplished using vertical drill holes. However in some
          mines this has lead to shallow slope failures on the newly formed highwalls, largely due to
          the presence of steeply dipping joint planes dipping into the highwall. When this occurs
          vertical pre-splitting cannot be used. A much improved result has been achieved by using an
          angle presplit. An angle of 70 degrees has often been employed in this application. Best
          results are obtained if the subsequent production blast is drilled on the same angle, so that a
          constant burden from crest to toe can be maintained on the front row. The weight of charge
          can be obtained by calculating the diameter required of a distributed decoupled charge of
          the explosive.




(iii) Cushion blasting - Cushion blasting is a common controlled blasting technique in surface
operations, second to pre-splitting as the most common method. Cushion blasting is used to slash or
trim excess material from the bench face to leave a smooth, clean wall with little backbreak, which
will remain stable for extended periods.

Blastholes are drilled in a line along the planned excavation limit and are loaded with a reduced
charge capable of slashing material from the wall without damaging the rock behind the holes.
Charges are usually decoupled for this purpose. Common diameters used in cushion blasting have
been 4-7 inches (102-178 mm), but large holes have often been used in open pit mines. In the
common range of diameters hole spacing of 5-8 feet (1.5-2.4 m) have been typical. In large
diameters hole spacing are greater. As a general rule the spacing in feet should be 1.25 to 2.0 times
the hole diameter in inches. The lower value is to be used in hard, competent rock while the higher
value applies to soft, highly fractured rock. Following figure illustrates a single row cushion blast
using 12¼-inch (311 mm) boreholes. A coupled toe charge is followed by a decoupled column
charge. Note the projected break line which will leave an angled face at the excavation limit. The
coupling ratio in this case is about 0.37 for the column charge. As an alternative to a decoupled
charge low density explosives could be used in a cushion row. Gassed slurries or emulsions are an

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example. The density of ANFO can be reduced by adding micro-balloons or perlite. As the density is
decreased the velocity of detonation and the borehole pressures also decrease.




Cushion blasting is also performed using multiple row blasts. These usually incorporate larger
diameter holes in the 9 7/8—12 ¼ -inch range (251—311 mm). However, in surface gold mining
diameters are more typically 6 3/4—7 7/8-inch (171—200 mm). These blasts typically consist of
three to four rows including the cushion row. This type of final wall blast is typically called a trim
shot and the cushion line is then termed the trim row. These blasts consist of three components
similar to a presplit blast. (a) The trim row, (b) The buffer row and (c) One or more production rows.
The trim row should be suitably decoupled. The coupling ratio typically does not exceed 0.45.
Decoupling is often achieved using undersized cardboard tubes. An alternative is to use undersized
plastic liners manufactured for use in pre-splitting and trim blasting. A third approach is to place a
suitable charge in the bottom of the hole and allow the gases to expand into the void above. The
trim row must do sufficient work on the surrounding rock to slash excess material off the wall.
Therefore, borehole pressures greater than that required for pre-splitting are necessary and these
need to be sustained for longer periods. Thus when a concentrated charge is loaded in the bottom of
the hole the use of an airbag and stemming may be a good way to contain the explosion gases for a
longer time while still allowing the borehole pressures to be attenuated.

As with pre-splitting the last row of the production blast is a buffer row. The design is essentially the
same as is used in pre-splitting. A greater scaled depth of burial is achieved by increasing the
stemming thereby avoiding cratering back through the trim row at the crest. Blast timing for both
the shots is standard for the production and buffer rows and could be a V-1 tie-in for a square
pattern or a V-2 arrangement for a staggered square or equilateral pattern. The trim row should
detonate one delay period after the adjacent production holes.



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Two or more trim holes can be shot per delay provided these do not outrun the production holes
and any vibration considerations are accounted for. Good relief for the trim blasts to move away
from the final wall is essential. Firing across two free faces will be very useful. Adequate delay time
should be provided to allow for good relief.
For cushion blasting in general accurate drilling is required. Coupling ratio of 0.45 or less should
normally be used.

(iv) Line Drilling - This method is seldom used in open pit mines because the closely spaced holes are
costly. However, it has been used in some cases where the rock was very weak and difficult to
presplit or cushion blast. It is more commonly used in civil construction projects where overbreak
can be very costly.

The typical hole sizes for line drilling are 21½ to 3-inches (64-76 mm). However, large diameter
rotary drill holes can also be used. When the spacing between the holes remains constant regardless
of hole diameter the cost is comparable in small and large diameter work. If the spacing can be
increased as larger holes are used, then the larger diameters will be more economical. In small
diameter work hole depths should be restricted to 30 to 40 feet (9.1-12.2 m) to minimize hole
wander. Greater depths are possible when larger diameters are used. No sub-grade drilling is
needed. Drilling must be very accurate for line drilling to be successful. The holes must be drilled so
that they all lie in one plane which corresponds to the angle of the final pit wall. Unequal spacing
between holes will lead to variable results. A buffer row is once again essential to good results. The
design of the buffer row would be as discussed earlier.

(v) Air deck-Air shock techniques - Air-decking is a method which involves the use of a concentrated
charge in the blasthole with a void above the explosive. The idea was originally expounded by
Melnikov in 1940 but widespread use of the technique only developed during the 1980's. It has been
used in pre-splitting where a charge is placed in the bottom of the hole and an air-bag is placed near
the top of the hole with stemming above.

The gases from detonation freely expand into the void and the pressure is attenuated as would be
the case with a distributed charge or a concentrated charge when no stemming is used. However,
the explosion gases are contained in the blasthole for a longer period of time, due to the stemming,
and exert pressure on the borehole wall for a longer time. Thus the stress generated in the ground
between holes is sustained for more time and there is greater potential for wedging action to further
open the presplit crack.

Experience in the industry has been that the explosive consumption should be 8 to 11 percent of the
total blasthole volume and 14 to 18 percent with respect to the air-deck volume above the charge.
However, one should also check the decoupled borehole pressures using the methods above to
insure that these pressures will suit the rock being presplit. When an air-bag and stemming are used
it may be possible to increase the hole spacing. However, this needs to be assessed on a site-by-site
basis. Geology will play an important role in determining whether spacing can be expanded beyond
those used in conventional techniques.


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The best approach will be to initially design the presplit shot, using the air-deck approach, on the
normal presplit spacing. If the results are of high quality increase the spacing by 20 percent. If good
results are still obtained increase the spacing in 10 percent increments until the optimum is
achieved. In the final analysis geology is likely to be the determining factor for the success of air-
decking on the presplit row. The method has been used to good effect in strata with horizontal
bedding that is relatively widely spaced.

Moreover, highly fractured rock tends to lead to a poorer result. Containing the decoupled borehole
pressure for a longer time can loosen existing joints and fractures as well as further defining a
presplit crack. The hole spacing is more likely to be controlled by the distance between major joints
than by the application of air-deck technology.

Often in mining and construction, blasting takes place in proximity to housing and other unowned
structures. Under these circumstances presplit holes cannot be left un-stemmed because the
resulting airblast becomes excessive. In larger diameters the use of an airbag allows the hole to be
sealed such that it can be stemmed and airblast reduced to acceptable levels. In small diameters the
hole may be plugged by simply pushing a wad of plastic hole liner down to the desired depth. Air-
decking technology may have good application on the buffer row as well. A bulk loaded charge could
be placed in the blasthole with an air-deck above and then stemming above the inflated air-bag. In
this manner the borehole pressure could be reduced while being distributed evenly throughout the
hole. Crushing around the hole and crest fracture can be avoided provided the plug is placed at the
correct depth. When active highwall pre-splitting is employed in deep holes the weight of explosive
needed to provide a suitable decoupled borehole pressure can become large. This can lead to
excessive fracturing around the toe of the hole. Thus there could be an advantage to splitting the
charge into two and placing these at different locations in the hole to reduce the potential damage.
An air-bag could be placed at the appropriate location and the upper charge placed above it thereby
reducing the potential for damage.

Air-decking technology may have good application on the buffer row. A bulk loaded charge could be
placed in the blasthole with an air-deck above and then stemming above the inflated air-bag. In this
manner the borehole pressure could be reduced while being distributed evenly throughout the hole.
Crushing around the hole and crest fracture can be avoided provided the plug is placed at the
correct depth. When active highwall pre-splitting is employed in deep holes the weight of explosive
needed to provide a suitable decoupled borehole pressure can become large. This can lead to
excessive fracturing around the toe of the hole. Thus there could be an advantage to splitting the
charge into two and placing these at different locations in the hole to reduce the potential damage.
An air-bag could be placed at the appropriate location and the upper charge placed above it thereby
reducing the potential for damage.

b. Blast Design for Final Wall Shots - Successful wall control blasting involves not only the wall
control row and the buffer row but also the design of the associated production blasts. If the overall
design is improper results will be poor, even though the wall control row has been well designed. A
key to successful wall control blasting is to allow excellent relief for the blast to pull away from the


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excavation limit. Achieving this result is a function of the orientation and millisecond delay timing of
the shot.

When two free faces are available the blast is better able to pull away from the final wall. The shot
can be delayed to systematically pull the buffer row holes away from the presplit or trim line one
hole at a time. The potential for freezing material to the face or wall damage is greatly reduced.
When tieing-in the blast the orientation can be V-1 at 45° to the free face if the pattern is square. If
it is a staggered square or staggered equilateral pattern the shot may be tied-in on the V-2
orientation along the long axis at a 34 degree or 30 degree angle to the principal free face
respectively. These latter patterns have often given good results, based on the substantial burden
reduction across the tie-in lines and the consequent ability to displace the material away from the
wall. If only one free face is available then a full echelon tie-in can be used, oriented to the single
free face.

Blast Damage Mechanisms: There are a number of basic principles that can be applied for limiting blast
damage, but first, we need to understand the mechanism of blast damage along final walls. There are a few
possible mechanisms that need to be considered, these being damage due to gas energy penetration into pre-
existing crack systems in the rock behind the blast, vibration related damage and geometrical effects. These are
briefly summarised below.
     • Damage by gas penetration - Originally, it was believed that blast damage was caused mostly by
          explosion gases entering planar weaknesses in the rock and forcing them open. However, research
          work carried out by Brent and Smith (1999) illustrated that the pressures in the rock behind a blast,
          even as close as one burden, are negative and not positive. In other words, damage in the final wall is
          unlikely to be the result of gas pressure penetration into a pre-existing network of joints, bedding
          planes and faults.
     • Damage caused by high vibration amplitudes - At the same time, Rorke and Milev presented
          information on rock damage as a function of vibration amplitude. Their measurements indicated that
          fresh cracks (damage) in quartzite occurred at amplitudes above 650 mm/s. Therefore, in this case, we
          can consider vibration amplitude as being a primary driver of rock damage. The variables that affect
          vibration amplitude are:
          i. Hole diameter
          ii. Charge mass per delay
          iii. Firing delays and sequence of firing
          iv. Firing time accuracy
          v. Level of confinement (burden)
          vi. The presence or absence of air decks
          These factors can be applied to alter the predicted vibration generated by the back row of holes in a
          blast. Predicting near field vibration requires a different attenuation model than the standard scaled
          distance equation. In normal vibration prediction, the source (such as a blasthole or a blast) can be
          regarded as a single point source because it is far from the point of concern. However, close to
          blastholes, predicting vibration is a little more comple, as each individual element of charge
          contributes to the vibration as a point charge.
     • Damage related to pit-wall and blast geometry - Very often, pit-wall and blast geometry are ignored.
          However, they can be a major source of unwanted wall damage. The width, length and height of the
          trim blast have an impact. The angle of the faces along the final wall (presplit plane angle if pre-
          splitting is being done) and whether double or single bench pre-splitting is applied will also influence
          the damage results.
Therefore, during the design stages of the final pit geometry, decisions about the presplit angles and presplit
heights should be made by considering the nature of blast related damage that can occur as a result of poor
choices.



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The millisecond delay timing must be sufficient to allow the rock mass to displace freely. Delay times
of 2 to 3 times the effective burden on the tie-in should be considered minimum. In some quarry
applications delay times of 5 to 7 times the effective burden have proven most effective. In weak
overburden (<5000 psi uniaxial compressive strength) above a coal seam a time of 6 ms per foot
(19.7 ms/meter) of effective burden has resulted in the ability to pull the casting shot cleanly away
from the active highwall presplit leaving a smooth wall and a good cost profile for subsequent
operations. In row on row casting shots the delay time per foot (meter) of effective burden may
vary, being least on the front rows and increasing further back in the pattern to create more relief.

5. WALL CONTROL PRACTICES IN UNDERGROUND OPERATIONS:
Controlling overbreak is important in underground mining and tunneling. Control of the blast effects
at the perimeter can reduce the amount of support needed in drifts and tunnels. Equipment and
facilities can be more readily installed.




In stopes leaving a smooth wall contributes to safety. Ore dilution is minimized when overbreak is
minimized which can have a major impact on mining costs. For example, one study in VCR stopes has
shown that controlled blasting reduced dilution from 20-35 percent to 3-9 percent (Plis, et al, 1991).

Many of the principles discussed above are applicable to wall control blasting underground. The goal
at the perimeter is to reduce the explosive loadings and the borehole pressures in order to avoid
damage to the wall and minimize the overbreak. As before the degree of decoupling needed to
provide a desired borehole pressure can be calculated. Coupled borehole pressures can usually be
obtained from the manufacturer for a given explosive; otherwise may be calculated.

Decoupling, to reduce the pressure, can be achieved by employing undersized cartridges of a
suitable product. Given hole diameters used underground and the diameter range of typical
explosives a coupling ratio in the range of 0.4 to 0.5 is common. Added decoupling may be achieved
if a space is left between cartridges as the explosive is loaded. In some cases a coupled, low density
product is used. The velocity of detonation and hence the borehole pressure is reduced without the



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inconvenience of decoupled loading. Low density products can include ANFO (with micro-balloons or
polystyrene) slurries and emulsion.

Another approach to reducing the pressure is to trace the borehole with a high grain count
detonating cord or a cord and a very light powder load. Pressures can be varied depending on the
grain count of the cord used. The primary difficulty with this approach is that it can be difficult to
lock the product in the blasthole and high grain count products may not meet underground fume
class regulations. The row of holes next the perimeter holes (often called the cushion row) must be
carefully designed. This is necessary to avoid damage beyond the perimeter of the excavation. These
holes are sometimes loaded as production holes rather than as buffer holes. However, it is often
appropriate to adjust the charge weights in these boreholes which may be done by not tamping the
explosive or by decoupling.

Explosive loading in smooth-wall blasting typically ranges from 0.10 lb/ft2 to 0.20 lb/ft2 (0.49 kg/m2
to 0.98 kg/m2). The actual load will depend on the rock strength, and the degree of weathering or
fracturing experienced. To obtain a good result with smooth-wall blasting it is essential that the
boreholes be drilled parallel. Varying spacing between holes will lead to poor results, just as in
surface operations. Inaccurately drilled holes next to the smooth-wall holes will lead to damage into
the wall or poorly fragmented material at the back and sides. Therefore, suitable procedures for
layout and drilling of the blastholes are a prerequisite for good results. Following figure shows a
typical smooth-wall blast design:




Overbreak in stoping operations is both a safety problem and may also result in excessive dilution.
These problems have become more acute as many mines have adopted vertical crater retreat

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methods in large diameter holes. Considerably more energy can be concentrated next the
hangingwall in a 6 1/2—7 7/8-inch (165 to 200 mm) hole than in a 2-inch (51 mm) hole. Therefore
appropriate precautions, involving controlled blasting techniques, must be taken to control the
breakage at the walls of the stope. Once again the requirement is to reduce the energy at the
perimeter, especially the hanging-wall, through decoupling and reduced explosive loading. The rock
is slashed away from the wall without resulting in damage to the perimeter that results in overbreak.

As is true in drifts and tunnels and surface operations as well, accurate drilling is essential to good
wall control results in stopes. Even if everything else is done correctly the wall control will be poor if
the bore holes have not been correctly drilled. Another factor essential to success is that the blast be
shot to good relief. This requirement holds true for all controlled blasting work underground and on
surface as well. Since most underground blasts are quite confined by nature particular attention
must be paid to how the shot is opened and how it is timed, to maximize relief without disrupting
holes firing on subsequent delays.

In longhole stopes employing small diameters similar principles apply. Borehole pressures along the
hangingwall should be reduced. The amount of reduction should be keyed to the rock strength and
geological structure as discussed above. Accurate drilling is essential and the delay timing pattern
should be designed for maximum relief away from the wall. In stopes, as in drifts and tunnels, the
most common approach is to use a smoothwall technique. Therefore the wall control holes are
detonated last for the purpose of slashing the remaining material off the perimeter leaving a
competent wall at the designed excavation limit.

In summary the following items are very important in underground wall control blasting:
       • Parallelism and accuracy in drilling must be maintained on the perimeter holes. Indeed all
       holes in the round must be accurately drilled for best results.
       • Blasting energy must be reduced at the perimeter through explosive selection and/or
       decoupling.
       • Loading factors typically range from 0.1 to 0.2 lbs/ft2 (0.49 to 0.98 kg/m2). Actual loads
       depend on rock strength and fracturing.
       • The collars of the wall control holes must be plugged to prevent ejection of the explosive.
       • Spacing between wall control holes will normally be reduced compared to that of
       production holes.
       • Proper location of the production holes next the perimeter holes are essential to avoid
       breaking beyond the limit.
       • In drifts and tunnels drilling to an arch rather than a flat back may improve results.
       • Wall control holes are normally fired on the last delay in the manner of cushion blasting to
       slash the remaining material off the wall.
       • The blast should be opened and timed for maximum relief at the perimeter.

6. Controlled Blasting on Construction Projects:
The principles already expounded cover the majority of controlled blasting requirements on
construction work as well. Primary needs are to control the energy and borehole pressures at the
limit of the excavation through decoupling or, possibly, the use of low density explosives. The buffer

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row must be properly loaded and located relative to the wall control line. The shot should
incorporate the maximum ability to relieve away from the perimeter and should be delayed to
provide excellent relief for the next holes to detonate.

In surface construction line drilling is more often employed than in mining or quarrying operations. It
may be the best method for very close in blasting when vibration levels associated with pre-splitting
would be unacceptable. Where highly accurate results must be obtained line drilling using very
closely spaced holes, while expensive, can provide the best result.

IMPROVEMENT OF SLOPE STABILITY: The quarry operator should plan for and mitigate the blast damage to
the final walls by implementing a proper production blast design and employing controlled blasting techniques.
Production blasting should be designed to limit rock fracturing behind final wall. In addition, controlled blasting
techniques such as pre-shearing (presplit) and cushioning blasting techniques should be employed to define the
final face.

Production Blasting- To improve slope stability and avoid overbreak or backbreak (rock volume broken beyond
the plane defined by the last row of blast holes) during production blasting, the following precautions should be
addressed in the production blasting design by the quarry operator:

1. Avoid choke blasting into excessive burden or broken muck piles. Choke blasting is blasting with insufficient
expansion space and is considered a misfire.
2. Design the front row of blast holes to account for and move the burden. Burden is the distance between the
free face and the first line of blast holes.

3. Design stemming in the hole to account for the burden, diameter of the blast hole, and the unconfined
compressive strength of the rock. Stemming is inert material such as gravel inserted into the collar of the drill
hole to confine the explosive gases.

4. Use adequate delays and timing intervals for movement of the rock to the free face and creation of
additional free faces for the blast holes behind the present free face. The ratio of the timing between shot rows
to the burden should fall between 4 and 6 to minimize overbreak.

5. Employ delays between blast holes and rows to control the maximum instantaneous explosive charge.

6. Drill back row blast holes and “buffer holes” an optimum distance from the final face to facilitate excavation
and minimize damage to the final wall.

Controlled Blasting- The following precautions should be addressed in the controlled blasting design by the
quarry operator:

1. Employ pre-shearing only when the burden is adequate to contain the explosive energy along the shear line.
Burden should be equal or greater than the bench height.

2. Employ cushion blasting when the burden is inadequate to contain the explosive energy along the shear line.
For example when the burden is less than the bench height.

3. Design for closely jointed and sheared rock because highly confined gasses produced from the pre-shear
blast generated along the shear line may vent into the fractures causing damage to the rock face.

4. Employ test blasts for final design of controlled blasting.




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In surface work pre-splitting is the most common approach used. Buffer and cushion blasting are not
common for construction projects. Presplit holes are commonly drilled vertically in construction
applications. However, for projects such as a major road-cut that is to be presplit angled holes
should be considered. Moreover, construction work is often conducted in proximity to built up
areas. Therefore, airblast is an important concern. For this reason leaving presplit holes unstemmed
is not usually possible. Adequately stemming the boreholes prevents excessive airblast. In larger
holes airbags may be an appropriate way to plug the top of the hole for stemming.

Blast vibration can be an important issue on construction projects. Therefore, detonating a large
number of presplit holes instantaneously is often not possible. A delay will need to be introduced
into the line periodically. A shorter delay will be preferable and a 17 ms unit may be most
appropriate. When choosing the delay time to introduce into the presplit row some experimentation
may be appropriate to determine that delay duration gives the least vibration from the highly
confined presplit holes while still yielding a good presplit result.

When tunnelling or performing other construction work underground wall control blasting is very
much the same as that described for tunnelling and drifting underground. However, the need for a
good smooth-wall result can be even more critical. Irregular tunnel walls with considerable
overbreak can be very costly in terms of extra concrete requirements or other construction tasks
which become more difficult and time consuming.

7. CONCLUSION:
The blasting remains most inexpensive method of hard rock fragmentation, however, the cost
associated with blast damage in terms of safety and productivity of surface and underground
excavations is becoming increasingly important. Rock damage due to blasting is directly related to
the level of stress experienced by the rock and its pre-blasting condition. In high stress environment
and under unfavourable geological conditions, disturbances associated with blasting may result in
excessive ground control and dilution problems. To minimize these undesirable effects, perimeter
control techniques are available, but results of their application are often less than optimal. Critical
evaluation of factors influencing blast damage is to be studied properly for each case, in order to
understand well the nature and extent of rock damage caused by blasting. The factors influencing
blast damage can be broadly categorized in three areas: (i) Rock-mass features, (ii) Explosive
characteristics and distribution and (iii) Blast design and execution; rock-mass features cannot be
changed but their knowledge facilitates the judicious selection of explosive characteristics and blast
design parameters to obtain optimal results.

Thus, wall control blasting practices are necessary to reduce the impact of blasting on mine faces but
can also have a significant negative impact on mine productivity and operating costs. To provide
burden relief the trim blasts have fewer rows than full production blasts and are fired to a cleared
free-face; hence they are termed 'unchoked’. This practice leads to scheduling constraints on the pit
operations and can cause ore dilution due to excessive muckpile movement. The use of such trim
blasts stems from the perception that increased wall damage results from 'choked' blasts. These
concerns are based on the assumptions that blast vibration levels and explosive gas penetration
increase with increased blast burden and face confinement.

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References:
* Bauer, G.F. and Donaldson, D.M.; Perimeter Control in Development and Breasting By Use of a Blasting
                                                   th
Program Readily Accepted By Miners; Proc. of the 18 Annual Conference on Explosives and Blasting
Technique; January, 1992; Orlando, Florida.

* Calder, P.N.; Pit Slope Manual, Chapter 7-Perimeter Blasting; CANMET; Report 77-14; May, 1977.

                                                                                              th
* Calder, P.N. and Tuomi, J.N.; Control Blasting at Sherman Mine; Proceedings of the 6 Conference on
Explosives and Blasting Technique; ISEE; 1980.

* Chiappetta, F. and Mammele, M.; Analytical High Speed Photography to Evaluate Air Decks, Stemming
                                                                                           nd
Retention and Gas Confinement in Presplitting, Reclamation and Gross Motion Applications; 2 Int.
Symposium on Rock Fragmentation by Blasting; Keystone, Colorado, 1987.

* Crosby, William A. and Bauer, Alan; Wall Control Blasting in Open Pits; Mining Engineering; February, 1982;
pp 155-158.

* Hunter, Christopher, Fedak, K. and Todoaschuck, J.; Development of Low Density Explosives with Wall
Control Applications; Nineteenth Annual Conference on Explosive and Blasting Technique; ISEE; January, 1993;
San Diego, CA.

                                                             th
* Livingston, C.W.; Theory of Fragmentation in Blasting; 6        Drilling and Blasting Symposium, University of
Minnesota, 1956.

* Plis, Matthew; Fletcher, Larry; Stachura, Virgil; Sterk, Paul; Overbreak Control in VCR Stopes at the
                                         th
Homestake Mine; Proceedings of the 17 Conference on Explosives and Blasting Technique; ISEE, February,
1991, Las Vegas, Nevada.

* Pilshaw, Russel N.; Rock Products; 1991.

                                                                                         th
* Revey, Gordon F.; Controlled Blasting at the Hanging Lake Tunnels; Proc. of the 17 Annual Conference on
Explosives and Blasting Technique; ISEE, February, 1991; Las Vegas, Nevada.

* Bhandari, S., 1997, Engineering rock blasting operations, A.A.Balkema, Rotterdam.

* Singh, P.K., Roy, M.P., Joshi, A., Joshi, V.P., 2009, Controlled blasting (pre-splitting) at an open-pit mine in
India, Proc. Int. symposium on “Rock fragmentation by blasting”, Fragblast9, Granada (Spain), pp 481-489.

* Partha Das Sharma; ‘Controlled Blasting Techniques – Means to mitigate adverse impact of blasting’; Procc.
    nd                                                                  th   th
of 2 Asian Mining Congress, organized by MGMI at Kolkata (India) dt. 17 – 19 January 2008 (pp: 286 –
295).

* Partha Das Sharma; “Application of Air-Deck Technique in Surface Blasting”,
(http://miningandblasting.wordpress.com/2010/06/26/application-of-air-deck-technique-in-surface-blasting/ )

* Duncan C. Wyllie, Christopher W. Mah; “Rock slope engineering: civil and mining”, Inst. of Mining and
Metallurgy.



                Author: Partha Das Sharma, (B.Tech-Hons., Mining Engg.),
E.mail: sharmapd1@gmail.com, Website: http://miningandblasting.wordpress.com/ Page 23
WALL CONTROL BLASTING TECHNIQUES
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* Konya, C.J., 2003, Blast Design in Rock Blasting and Overbreak Control, 2nd Edition, National Highway
Institute, Federal Highway Administration, pp. 226-287.

* Wyllie, Duncan, C., Mah, C.W., 2004, Rock Slope Engineering: Civil and Mining, 4th Edition, Spon Press, pp.
245-275
* Brent, G.F. and Smith, G.E. 1999, The detection of blast damage by borehole pressure measurement.
Fragblast 6, Johannesburg, South African Institute of Mining and Metallurgy, pp 9 – 13.
* Rorke, A.J. and Milev, A.M., Near field vibration monitoring and associated rock damage. Fragblast 6,
Johannesburg, South African Institute of Mining and Metallurgy, pp 19 - 22.
* Holmberg, R., and P-A. Persson. 1978. The Swedish approach to contour blasting. Proceedings of the 4th
Conference on Explosives and Blasting Technique. SEE. Pp 113-127.
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Author’s Bio-data:




Partha Das Sharma is Graduate (B.Tech – Hons.) in Mining Engineering from IIT, Kharagpur, India (1979) and
was associated with number of mining and explosives organizations, namely MOIL, BALCO, Century Cement,
Anil Chemicals, VBC Industries, Maharashtra Explosives and Solar Expl., before being a Consultant.

Author has presented number of technical papers in many of the seminars and journals on varied topics like
Overburden side casting by blasting, Blast induced Ground Vibration and its control, Tunnel blasting, Drilling &
blasting in metalliferous underground mines, Controlled blasting techniques, Development of Non-primary
explosive detonators (NPED), Hot hole blasting, Signature hole blast analysis with Electronic detonator etc.
Author’s Published Books:
1. "Acid mine drainage (AMD) and It's control", Lambert Academic Publishing, Germany,
(ISBN 978-3-8383-5522-1).
2. “Mining and Blasting Techniques”, LAP Lambert Academic Publishing, Germany,
(ISBN 978-3-8383-7439-0).
3. “Mining Operations”, LAP Lambert Academic Publishing, Germany,
(ISBN: 978-3-8383-8172-5).

Currently, author has following useful blogs on Web:
http://miningandblasting.wordpress.com/
http://saferenvironment.wordpress.com
http://www.environmentengineering.blogspot.com
www.coalandfuel.blogspot.com

Author can be contacted at E-mail: sharmapd1@gmail.com, sharmapd1@rediffmail.com,
-------------------------------------------------------------------------------------------------------------------
Disclaimer: Views expressed in the article are solely of the author’s own and do not necessarily belong to any
of the Company.
                                                                ***

                Author: Partha Das Sharma, (B.Tech-Hons., Mining Engg.),
E.mail: sharmapd1@gmail.com, Website: http://miningandblasting.wordpress.com/ Page 24

				
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Description: Wall failures are costly and often life threatening. The goal of efficient wall control blasting is to make transition from a well fragmented rock mass to an undamaged slope in as short a distance as possible. This can be quite challenging due to the many factors that influence wall damage. To develop efficient designs one must have a basic understanding of wall failure mechanisms as well as limitations of wall control procedures.