wall control blasting techniques

Author: Partha Das Sharma, (B.Tech-Hons., Mining Engg.),

E.mail: [email protected], Website: http://miningandblasting.wordpress.com/

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

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

E.mail: [email protected],

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.


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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.


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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 ρρρρ 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.


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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.

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

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.


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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.


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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


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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


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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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• 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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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/m

2). 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

Program Readily Accepted By Miners; Proc. of the 18th

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.

* Calder, P.N. and Tuomi, J.N.; Control Blasting at Sherman Mine; Proceedings of the 6th

Conference on

Explosives and Blasting Technique; ISEE; 1980.

* Chiappetta, F. and Mammele, M.; Analytical High Speed Photography to Evaluate Air Decks, Stemming

Retention and Gas Confinement in Presplitting, Reclamation and Gross Motion Applications; 2nd

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.

* Livingston, C.W.; Theory of Fragmentation in Blasting; 6th

Drilling and Blasting Symposium, University of

Minnesota, 1956.

* Plis, Matthew; Fletcher, Larry; Stachura, Virgil; Sterk, Paul; Overbreak Control in VCR Stopes at the

Homestake Mine; Proceedings of the 17th

Conference on Explosives and Blasting Technique; ISEE, February,

1991, Las Vegas, Nevada.

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

* Revey, Gordon F.; Controlled Blasting at the Hanging Lake Tunnels; Proc. of the 17th

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.

of 2nd

Asian Mining Congress, organized by MGMI at Kolkata (India) dt. 17th

– 19th

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.


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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: [email protected], [email protected],

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Disclaimer: Views expressed in the article are solely of the author’s own and do not necessarily belong to any

of the Company.

***

wall control blasting techniques

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