Read this article to learn about the blasting techniques adopted in underground and opencast mines.
Blasting Techniques Adopted in Underground Mines:
Drilling Patterns in Stone:
Drilling patterns, also called shot-hole patterns, are named after the type of cut holes used and the principal patterns are:
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(1) Pyramid cut or cone cut
(2) Wedge cut
(3) Drage cut
(4) Fan cut
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(5) Burn cut
(6) Coromant cut
(7) Ring drilling.
In general each hole in a round covers an area of 0.3 to 0.5; cut holes are located about 0.45 m vertically apart, and easers 0.5 to 0.6 m apart; trimmers are drilled at intervals of about 0.6 m round the perimeter of the drift.
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(1) Pyramid Cut or Cone Cut:
Pyramid cut consists in drilling holes (in the centre of the drift axis) at corners of a square, 0.7 to 1 m sides, almost to meet at a point at the back of the round. In a modified design known as cone cut, illustrated in Fig. 8.6, holes are drilled forming corners of a polygon with centre holes, all nearly meeting at a point at the back of the cut. The depth of pyramid cut is generally restricted to 50% to 60% of the width of the drift.
(2) Wedge Cut:
Wedge cut takes the form illustrated in Fig. 8.6 in which 2 to 4 pairs of holes are drilled to form a wedge, each pair starting from two sides of the drift centre and inclined at an angle less than 45° towards the centre almost meeting at the back of the cut along a line.
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Pyramid and wedge cuts, the most commonly used forms, are suitable for uniform, thickly bedded and hard rocks. They consume the least total quantity of explosives, but the depth of pull is restricted by the width of the drift.
(3) Drag Cut:
Drag cut, used for small drifts, 1.8 to 2.4 m wide, consist in drilling holes at an angle to the cleavage so that strata break along the cleavage planes (Fig. 8.6). This pattern, being dependent on the direction of cleavage planes, calls for frequent changes, which are detrimental to systematic work and the pattern is, therefore, not favoured for large excavations.
(4) Fan Cut:
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Fan cut, favoured for laminated strata, mostly soft, covers the face with a fanlike pattern as shown in Fig. 8.6 As each shot has to act for itself, charge in each hole is heavy. This cut is not recommended for hard ground.
With the cuts it is normally difficult to drill deeper than half the width of the drift because of the angle of the drill, but with the burn cut, advance equal to the width of the drift can be obtained.
(5) Burn Cut:
In Burn cut, parallel holes at right angles to the face are drilled in a cluster which may take several forms. Some of the holes (which are sometimes of larger dia. than the rest) are left uncharged to give relief to the heavy concentration of explosives in charged holes. This cut is effective in hard, brittle, homogeneous ground which breaks evenly; but cannot be used in springy plastic ground.
The advantages claimed for this cut are:
i. Drilling time is considerably reduced and supervision in drilling is easier as holes are straight.
ii. Depth of pull is dependent of the size of the drift, and
iii. The quantity of blasted material is not projected far with a suitable form of the cluster.
All the holes, charged as well as uncharged ones, may be drilled by the same drill and those to be left uncharged may be reamed by a reaming bit. At Mosaboni mine shot-holes for burn cut in the drives are 32 mm dia. and the central holes which is not charged is reamed to 58 mm. dia.
(6) The Coromant Cut:
The coromant cut is a new type of parallel hole cut which has been worked out with the object of achieving greater advance per round in tunnels or drifts of small area. In principle, the Coromant cut consists of a slot, which is left unloaded, together with 6 outer cut holes, the locations of which are carefully calculated.
All the drilling is done with the same pusher-leg drill. The slot is produced by the contiguous drilling of two holes. Drilling is carried out using a 20 mm drill rod with taper and a special drill bit of 57 mm diameter.
The hole is drilled first, being directed slightly upwards to assist in the disposal of cuttings. Continuous drilling calls for guidance in drilling the second hole. For this purpose a guide tube is used, being inserted into the first hole and secured by means of an expander.
Burn cut and coromant cut are known as parallel-hole cuts. In drilling parallel-hole cuts, precision is always an important factor. A special drill jig or template which gives the desired precision and facilities drilling, has, therefore been designed for the outer cut holes.
This template also drilling considerably faster. The template normally used is the 6-hole template which fits in with the standard drilling pattern of the Coronmant Cut. The two centre tubes of the template are slid into the finished slot, and the template is then secured in place by means of an expander in one of the tubes. The drilling of peripheral holes in the cut is then carried out using ordinary integral steels.
It may be possible to have only one centre hole depending on the nature of the rock. The charge varies with the dia. of the empty hole and in the case of the pattern shown, the recommended charge is 0.3-0.4 kg/m length of the hole. The holes are fired in a sequence shown by delay number in the pattern.
The cut holes and the rest of the holes in the drift are normally blasted in the same round with the help of mile second delay detonators. The standard coromant cut pattern covering six cut holes has been found to be good for the great majority of rocks. The arrangements of the holes and the firing order have been so chosen that the dislodged rock is given sufficient room to expand.
(7) Ring Drilling:
For some types of stopping operations in metalliferous mines drilling with long holes is practiced.
There are 2 types of ring drilling:
(a) Vertically ring drilling where rings are drilled in vertical planes, radially like a fan to break to a vertical face; applicable in sub-level stopping.
(b) Horizontal ring drilling where the holes are fanned out radially in horizontal planes to break to a horizontal face; applicable in shrinkage stopping.
The principle is the same as for vertical drilling and blasting. Short delay interval of 25 millisecond is usually employed between holes in each row or ring, starting from the easiest breaking section in the middle and progressing towards the walls.
Long hole blasting is done in wide or bodies with strong walls so that dilution is minimal. In narrow ore bodies, on the other hand, long hole blasting is not generally economical due to the high cost of drilling. The method is also not quite applicable in irregular ore bodies because long holes would either tend to leave ore unblasted in places or to break into waste rocks.
Solid Blasting:
In a development gallery, coal can be blasted without giving an undercut, by the use of explosives of P5 type. The blasting is known as blasting-off-the-solid or, in short solid blasting.
P5 explosive is used along with Carrick short delay detonators (of I.E.L) or coal delay detonators (of l.D.L.) Introduction of such blasting in mines needs approval from the D.G.M.S. for exemption under the following regulations:
(a) Reg. 173- To blast off the solid.
(b) Reg. 175- To use delay detonators.
(c) Reg. 168- (5)- To fire rounds of more than 6 shots in coal.
The pattern of shot-holes drilled for solid blasting in coal in galleries of bord and pillar working is:
1. Wedge cut, or
2. Fan cut
Drivage of a narrow gallery in coal without an undercut can be compared with drivage of a drift in stone. Blasting out coal in the centre of the face in the wedge cut pattern gives free face for the remaining coal.
The maximum permissible charge to be used per hole therefore varies-when blasting out the central wedge with P5 explosives and when blasting out the remaining cola (after the central wedge is blasted out thereby providing free face) with P1/P3 explosives if these are used. On a long wall face, the shot holes are drilled at an angle of 45° to 60° to the face; a coal face 2.4 m high requires 3 rows of shot holes and distance of holes in the same in the same row should be nearly 1.2 m.
Powder factor with solid blasting is 1.5 to 2.5 te of coal per kg of explosive and 0.8 to 1.25 te of coal per detonator.
DGMS stipulations on maximum permissible charge in a shot-pole, delay interval, etc. in u/ g coal mines.
I. Explosives:
II. Delay Detonators:
(a) While using non-incendive delay detonators in ‘BOS’ applications the maximum delay period between the 1st & last shot in a degree I and II gassy coal seam will not exceed 150 millisecond.
(b) While using non-incendive delay detonators in ‘BOS’ applications, the maximum delay period between the 1st & last shot in degree III gassy coal seems will not exceed 100 millisecond.
(c) The delay period between 2 consecutive shots with different delay numbers will not exceed 60 millisecond.
III. Distance between 2 adjacent shots with different delay numbers will not come closer than 0.6 m at the explosive-charged ends.
The advantages of solid blasting are:
1. It eliminates the use of coal cutting machines. The particularly important in steep seams where flitting and control of coal cutting machine is difficult and risky.
2. Skid mounted coal cutters are fitted with the help of anchor pipes which can be used in seams of upto 2.2 m height. Moreover, the roof and floor of the seam to use such pipes should be hard. Where these working conditions do not exist the skid mounted coal-cutter cannot be used and solid blasting is an advantages.
3. Saving in capital expenditure on equipment at the face, like coal cutting machine, gate- and boxes, cables, etc.
Compared to the undercut faces, the amount of drilling on a solid blasting coal face is high but the provision of one or two extra drills at the face is not very costly, considering the cost of coal cutting machines, gate-and-boxes and medium tension cables of special construction for remote control operation.
4. If tubs arc used for transport of coal, track can be laid right up to the face in the absence of coal cutting machines.
5. When working a seam over a caved area, the floor is irregular and coal cutting machines cannot be used. Solid blasting offers definite advantages in such cases and in other cases where floor may be irregular for some other reasons.
For Deg. 1 gassy mines at least 284 m3 of air per min. shall be conducted in the ventilation connection out-bye of every face where solid blasting is to be practiced.
For Deg. 2 gassy mines at least 284 m3 of air/min. shall be conducted upto every face where solid blasting is to be carried out.
The results obtained by solid blasting in one mine in Hazaribagh area (C.C.L.) are given below for guidance:
Gradient of seam – 1 in 3.7
Category of gassiness – degree 1.
Working method – Bord & pillar development
Gallery size – 4.2 m wide x 2.4 m high.
Shot holes at a face – 1.4 m deep, wedge cut pattern total No. 12
Explosive charge – 500 g/hole; total charge 6 kg.
Pull per round – 1.06 m.
Coal yield/kg of explosive – 2.0 te (soligex)
Coal yield/detonator – 0.9 te
One driller and two helpers drill 60-70 holes in a shift of 8 hours; one shot firer, 2 explosive carriers and 1 stemming material carrier constitute the blasting crew and fire nearly 60 shots in a shift of 8 hours.
Limitations on Carriage of Explosives:
Explosives and detonators should not be carried in the same box. Not only this, one person should not carry explosives and detonators together though they may be in different boxes. An explosive carrier can carry only 1 case containing 5 kg of explosive but Director of Mines Safety may give relaxation. Explosive case should be numbered. Jeep or trailer is allowed to carry more explosives for opencast mines for deep holes.
The charge for blasting in a hole may consists of one or more cartridges. It is desirable to have the least number of separate cartridges as far as possible, commensurate with the work to be done.
One of the cartridges should have a detonator inserted into it. Such cartridge, equipped with a detonator, is called a “primer cartridge”.
If an electrical detonator is used to prepare in primer cartridge, open the cartridge at one end, make hole with a pricker of brass or wood, insert the detonator until it is completely buried in the explosive, put back the flap of the cartridge and hitch the leading wires around the cartridge to prevent the detonator being withdrawn accidentally during charging. Permitted slurry explosive can be primed from the sides by pricking a hole into it.
A cartridge of non-combustible stemming material is first pushed in the hole. The charge is then placed in the hole and the primer cartridge pushed last of all, so that the “business end” of the detonator points towards the main body of the charge.
This position of primer cartridge is termed “direct initiation”. With this position of the detonator, the strongest wave is directed towards the back of the hole and the chances of all the cartridges being properly exploded are a maximum. Direct initiation is best to prevent ignition of fire damp, reduces risk of blow-out shots, and gives maximum yield of coal having a free face.
When the detonator is at the back of the charges and the “business end” points towards the front of the hole, it is called “inverse initiation”. This is not practiced on coal faces, but is adopted.
(a) In “sumping” or “cut” shots in shafts and tunnels, and
(b) When using delay action detonators to fire a round of shots.
After the charge is placed in the hole, the shot-hole is stemmed with stemming material, keeping ends of the leading wires of detonators out of the hole. The stemming material should be compact but plastic, consisting of sand and clay in 3 : 1 proportion, and dried at the surface in the form of cartridges 150 mm to 200 mm long and 25-30 mm dia. The first 2 or 3 stemming cartridges near the charge should be tamped lightly by the stemming rod and the rest should be tamped hard.
Complete Procedure for Shot Firer:
The whole procedure which the shot firer has to follow to fire shots using electric detonators is outlined below:
1. Test shot holes for breaks. If a crack is found, the hole shall not be charged.
2. Test for gas from the hole and within 8 m of the hole.
3. See that the shot hole is 150 mm less than the depth of the cut, if a coal cutting machine is used.
4. Mark direction of the shot hole on the roof or side where practicable.
5. Charge the hole with explosive; insert the primer last of all Don’t force a primed cartridge into a shot hole of small size.
6. Stem the hole, first lightly and then hard, upto its mouth.
7. Spray stone dust or water within 18 m of the area.
8. Warn the workers to clear up and post helpers at suitable places 27 m away in approach roads to prevent workers inadvertently entering the area.
9. Lay out the shot firing cable.
10. Test again for gas within 18 m.
11. Couple the firing cable to detonator wires. If more than one shot is to be fired, all connections of detonator leads should be in series.
12. Take shelter.
13. Couple shot firing cable ends to the exploder.
14- Shout a warning again; ensure that workers have taken shelter, and fire the shots by a sharp twist of the exploder key. If the charge does not explode, try again with a sharp twist of the exploder key.
15. Allow the fumes and gases to clear.
16. Return to the shot hole, examine the roof, sides and timber supports and shout “all clear” for the workers to return to their work if the place is safe. Otherwise, have it dressed by dressers, and supported by timber men before workers enter it.
At the end of the shift the shot firer should write a report about the quantity of explosives blasted and the place of blasting.
Ignition of Gases:
The main danger from explosives in U/G coal mine is the ignition of fire damp. It may take place in the following ways-
1. By the Flame and Hot Gases – Though this is a common cause, because of “lag on ignition”, a flame, coming in contact with gas, does not ignite it instantaneously. If the flame is sufficiently cooled during the interval of lag on ignition, which is a fraction of a second, the gas will not ignite. If the heat energy generated during blasting of the explosive is effectively converted into mechanical work, the explosion products are not hot enough to ignite the gas.
2. By Incompletely Detonated Explosive – Such explosive may continue, to burn like an ordinary combustible material.
3. By incandescent particles coming out of the shot hole after blasting, if such particles come in contact with coal dust or gas.
4. By the compression wave of the blast which may compress the gases in the cracks connected with the shot hole and raise the temperature of the compressed gas to such an extent as to ignite it. Such 20-fold compression is known to be sufficient to ignite all inflammable mixtures of fire-damp and air.
The gas in the breaks and fissures connected with shot hole and leading to a fire damp pocket cannot be easily removed and the presence of breaks and fissures is one of the commonest causes in explosions of fire damp due to shot firing. If any break is found, the shot must not be fired.
A blown out shot is one which has not done any useful work of blasting coal, but has ejected itself out of the shot hole. An undercharged hole will result in blown out shot. Similarly in an overcharged hole where only part of the charge is utilized in doing the useful work of blasting cola the excess charge may still be burning when the coal is broken, and it has the same effect of igniting gas or coal dust as an undercharged blow out shot.
Common Causes of Accidents:
1. Not taking Proper Cover. This is the most common cause of personal injury due to explosives. It is essential that the shot-firer shall himself take adequate cover and see that all workmen in the vicinity of a shot are removed to a safe–place.
No place in direct line with a shot can be regarded as safe and every person should be protected by at least one right angled corner. All approaches to the danger zone should be guarded by sentries or otherwise so as to prevent anyone entering inadvertently.
2. Failing to warn persons in an adjoining’ place into which the blasted rock may be thrown, as is possible when two galleries are about to join and partition is thin.
3. Carelessness in handling detonators causing them to explode or to be lost in a mine.
4. Carelessness whilst charging a hole, e.g. tamping too forcibly in the neighbourhood of the detonator, or ramming the primer cartridge into a hole of insufficient diameter.
5. Firing a shot when persons are at the shot hole due to instructions being misunderstood, or lack of proper sentries.
6. Returning to the face too early after firing a round of shots, one of which is a “hang-fire” (i.e. a delayed ignition), or before authorised to do so by the shot-firer.
7. Dealing with misfired shots otherwise than in the prescribed manner.
Ignition of gas or coal dust is a major accident which may arise from the use of explosives if proper precautions, as laid down in the Regulations, are not taken.
Misfired Shot:
When a detonator fails to explode, or after exploding fails to blast the charge of the main explosive cartridge, it is known as “misfire”.
The reasons for misfire are as follows:
1. Defective firing exploder.
2. Defective detonator or bad quality explosive either due to bad manufacture or due to deterioration during storage.
3. Broken detonator leads, defective or broken shot firing cable, bad connections between exploder and cable, between cable and detonator leads or at other places.
4. Short circuit of the cable or detonator leads due to poor or broken insulation.
5. Where fuse blasting is done, it may be due to-
(i) Wet fuse,
(ii) Improper timing of fuse, so that blasting by one fuse may cut the fuse of another hole, and
(iii) The fuse being drawn out during stemming.
To guard against misfire, the cable should be checked before blasting, the exploder should be examined once every three months by competent persons, and only good quality explosives and detonators, not spoilt during storage, should be used.
It is better to have two single-core cables for shot-firing separated by a good distance, instead of a twin core cable, which may be liable to short circuit if the insulation is bad. During use all the connections should be carefully made. For important blasting work, the circuit should be tested for continuity blasting work, the circuit should be tested for continuity by a galvanometer.
Dealing with a Misfired Shot Underground:
If shots are fired with safety fuse, no person should enter the site of blasting for 30 minutes after firing. If elec. detonators are used, this time may be reduced to 5 minutes after cable is disconnected from the exploder.
All the entrances to the place should be fenced.
Another attempt be made to blast it by making proper connections of the cable. If the misfired shot does not explode, the shot should be dislodged by drilling another relieving hole at least 0.3 m away from the misfired hole and by blasting it. The new hole should be drilled in the presence of the shot-firer who knows the directions of the misfired shot hole, so that during drilling the drill bit does not touch the misfired charge.
After the relieving shot has blasted the rock, a careful search for misfired cartridges and detonators should be made in the presence of the shot-firer in the material brought down by the shot. If the misfired explosive is not traced, the material should be loaded in a separate tub, distinctly marked, for further search on the surface.
The misfired explosive and detonator, when traced should be destroyed on the surface.
Except in the case of coal mines where the Regulations currently in force prohibit the placing of a second charge in a misfired hole, the stemming may be sludged out with compressed air or water under pressure, and the hole re-primed and fired. The stemming should not, however, be scraped out.
In underground coal mine the yield of coal on an undercut face is nearly 5-8 te per kg of explosive and 1-2.5 te per detonator but with solid blasting the figures are 1.8 to 2.7 te per kg of explosive and 0.8 to 1.35 te per detonator.
Some of the slurry explosives of I.D.L. have produced the following results during experiments at a few coal mines on precut face (Pentadyne figures are for solid blasting).
Maximum Shots Fired By a Shot Firer:
The mining Regulations fix the following limits on the number of shots to be fired by a shot firer.
Underground Coal Mines:
Deg. 1 gassy mines – 50 shots with single-shot exploder and 100 shots with multi shot exploder.
Deg. 2 & 3 gassy mines – 40 shots with single-shot and 80 shots with multi shot exploders.
If a mining sirdar or overman is working as a shot fire-also and has more than 30 person working under his charge, he can fire not more than half the number stated.
60 shots with single-shot exploder or with ordinary detonators and 120 shots with multi-shot exploder.
The D G M S can put restrictions on the above or give relaxation.
Blasting in Quarries:
Oxygen, gas liquefies at – 183°C. A given volume of liquid oxygen, when gaseous, is equivalent to 840 times at N.T.P. i.e., it has as much oxygen as would be available from 4000 times its volume of atmospheric air.
If a combustible ingredient, made in the form of a cartridge, is soaked in liquid oxygen and then subjected to reaction takes place with such terrific speed that large volume of a gas is instantaneously released at high temperature so as to cause explosion. The velocity of detonation under suitable conditions of confinement can be faster than 5000 m/sec. This is the principle behind the use of liquid ‘oxygen (LOX) as explosive.
LOX is used for removal of overburden as well as mineral in the quarries of coal as well as metalliferious mines. But its use is prohibited in underground coal mines. LOX is marketed by Indian Oxygen Ltd. in cartridges of two types.
i. Small cartridges of dia. 25 mm to 90 mm
ii. Large cartridges of di# over 100 mm upto 210 mm.
For small as well as large cartridges there is a “standard” cartridge. For small group the standard cartridge is 38 mm dia x 300 mm long. The standard cartridge of large group is called “Full cartridge” and it is 190 mm dia x 600 mm long.
The other cartridges are specified in terms of volume of the standard cartridge. In blasting performance, a full cartridge of LOX according to LOX dealers and results in the field, is considered equivalent to 18 kg of conventional explosives.
For those areas where the demand of liquid oxygen is heavy, the Indian Oxygen Ltd. has established central depots equipped with storage tanks and other arrangements of preparing LOX cartridges to be supplied to consumers about an hour or two before charging into blast holes. Such depots are at Bermo, Kathara, Lohardaga, and centrally located centres in mining areas. Liquid oxygen is transported from factories in special vessels on trucks.
A LOX cartridge ready for blasting is prepared at the depots by soaking an absorbent cartridge in liquid oxygen. The basic ingredient of an absorbent cartridge is a cellulosic substance like crushed jute stalk or other agriculture product though other substances such as hydrocarbons or metallic powders are used to impart to the soaked cartridge the properties of an industrial explosive.
The absorbent cartridge for use in coal contains a special composition called “Loxite- C”, with a view to reduce the temperature of the gaseous products after the blasting an such cartridges for use in coal are called LOX-C. The absorbent cartridge consists of the above ingredients wrapped in paper covered by eloth.
Loxite factory at Ranchi manufactures absorbent cartridges which are kept in stock at the depots. LOX cartridges are not stored in colliery magazines, but are supplied by depots on 2 to 3 hours’ prior intimation of the exact requirement for blasting.
For small consumers within 300 km of oxygen factories liquid oxygen is supplied in special containers of 26 litres by train.
During transit some evaporation, 5 to 10%, does take place. The small consumers have to obtain lixite absorbent cartridges, soaking boxes and other equipment to prepare LOX cartridges on the spot. Only small cartridges are prepared at the quarries in this manner for use in holes drilled by jack hammers or wagon drills.
LOX cartridges are inflammable and the flow of gaseous oxygen emanating from a cartridge will cause smouldering material, glowing coals, and cigarette stubs to burst into flames. LOX should therefore be kept away from such burning or smouldering material.
Characteristics of LOX are not constant. It depends much on the time that has elapsed between removal from the soaking vessels and firing. It is therefore not possible to specify the weight strength of LOX cartridge. LOX does not require a booster for blasting.
The LOX cartridges should be used in the field without delay (within half an hour in the case of large cartridges)’ to prevents loss of absorbed liquid oxygen. Grease or oil should not come in contact with the cartridges at any stage.
If the oxygen of a LOX cartridge evaporates, the absorbent, cartridge can be used again to prepare LOX cartridge at the depot. The ‘Life’ of a LOX cartridge in open is about one hour i.e. its full blasting power is available when used within that period but diminishes gradually after that period.
The “Life” however, considerably increases in the confinement of the hole. The cartridge looses its oxygen by gradual evaporation if it comes in contact with water. For use in watery holes only H-type cartridges wrapped in a polythene bag before lowering in the hole should be used. The toe will be not be sufficiently loosened if LOX is used without such precautions in watery holes.
LOX can be fired with or without the help of a detonator. A hole charged with LOX can be fired with safety fuse alone like gunpowder. In small-hole blasting, detonators are not much used as safety fuse is economic. Holes deeper than 3 m are fired by a detonating fuse.
Blasting Techniques Adopted in Opencast Mines:
Efficient blasting should give such rock fragmentation as to eliminate the need for secondary blasting. The fragmented rock should be easily handled by the bucket of the shovel. For mineral which has to go to crushers, the size should not be more than the input size for primary crusher.
Quantity of Explosive Consumption:
The volume of rock broken by explosive is given by the following formula –
Rock blasted per hole (solid m3) = depth of hole x burden x spacing.
Rock blasted by a round of holes = depth x burden x spacing x number of holes in the round
The following table gives an idea of explosive consumption in quarries (O.C.G. or 80% special gelatine) for various rocks under Indian conditions (Solid m3).
Depth and Pattern of Blast Holes:
The depth of vertical blast holes in coal is generally equal to the height of the bench. In hard rock like sandstone, laterite, hematite, etc. the depth should be 0.5 to 1 m deeper below the level of bench floor. This loosens the toe. In any case, the hole should terminate in hard rock and not in the soft one; otherwise the force of explosive is wasted.
Burden is the minimum distance from the face to the blast hole (the term usually refers to burden at the top of the face).
Toe is the projection of the bottom of a face beyond the crest. Sometimes the bottom edge of the bench is referred to as toe.
In hard rocks like laterite and hematite the spacing and burden vary from 0.3 to 0.4 times the height of bench. In less hard rocks like coal and sandstone, the spacing and burden vary from 0.5 to 0.8 times the height of bench. The exact dimensions depend upon the hole diameter, type of explosive used, type of the rock, nature of rock consolidation and the angle of cleats or laminations with the blast hole.
Preparation of Primer Cartridge and Charging Blast Holes:
For blasting with detonating fuse the primer cartridge of a booster, O.C.G. and similar gelatinous or semi-gelatinous explosive is prepared by pricking a through hole in it, 10-15 cm below the top with a brass or aluminium pricker and threading detonating fuse through it.
A knot is made at the end of the fuse in contact with the cartridge to prevent the former from coming out. In the case of LOX, to prepare a primer cartridge no hole is pierced in the cartridge but the detonating fuse is coiled at the neck of the cartridge in the form of a loose knot so that some length of fuse is in intimate contact with the explosive along its length.
Aquadyne and Superdyne slurry explosives which are in the form of cartridges closed at both ends by strong rubber bands can be made as primer cartridges by tying a detonating fuse round the cartridge. The short length of fuse at the knot end can be inserted into a small hole priced in the cartridge by a pricker. The slurry is viscous enough not to flow out of the small hole.
The primer cartridge is always placed at the bottom of the hole in case of blasting with O.C.G. and similar explosives and also in case of LOX, ANFO, and slurry explosives. It is generally lowered by the detonating fuse tied to it. LOX cartridges have a loop of copper wire round the neck.
A self-detaching hook is attached to the copper wire and the cartridge lowered in the hole. After the cartridge rests in its place in the hole the self-detaching hook comes out of the copper wire loop when the self-detaching hook comes out of the copper wire loop when the rope is slack.
Cartridges of other explosives including ANFO or slurry explosive are dropped or lowered into the hole. ANFO mixture or slurry which is not in the form of cartridges is poured into the hole. In the case of ANFO or slurry explosive a few cartridges of booster or O.C.G. in the middle of the total explosive column or at the top is essential for good blasting performance. The quantity of O.C.G. or other explosive as booster in the total explosive column if ANFO or slurry explosive is used is nearly 15-20%.
Deck loading i.e., separating the explosives charges into sections by placing a column of stemming between groups of cartridges is a useful technique for obtaining better rock fragmentation especially where soft and hard rock are encountered in alternate layers. A primer cartridge is placed at the deck charge also and the detonating fuse of bottom-most primer cartridge is threaded through the primer cartridge of the deck charge.
The drill cuttings in overburden are used for stemming after moistening them with water. The stemming is packed by tampers or a wooden tamping dolly.
LOX cartridges, if they have to be used in watery holes, should be enclosed in polythene bags before lowering. After LOX cartridges are lowered in the hole, sand bags are slipped on top of them to anchor them at the bottom of the hole and to prevent them from floating to the surface as liquid oxygen evaporates.
The hole is then completely stemmed by overburden cuttings. Gas bubbling through the water should not distract attention as the blasting efficiency is not affected if blasting is performed within a reasonable time.
The detonating fuse is cut after leaving nearly 1 m length outside the hole for attachment to truck line of fuse. The connections of detonating fuse of the blast hole with trunk line are shown in Fig. 8.21. Where two or more rows of blast holes are to be fired in one round, delay detonators are used.
Short delay detonators or millisecond detonators are used. Short delay detonators or millisecond detonators give good fragmentation and reduce ground vibrations, the latter advantage being important where quarries are situated near areas having buildings and costly engine foundations.
The blasted rock pieces in mechanised quarries fly upto 300 m. All the workers within 300 m of the blasting site should withdraw to safe places. Warning is given by the short firer half an hour before the blasting and again 5 minutes before the blasting by an electric siren, or alternatively by a bugle.
All the machinery like shovels, drills etc. should be withdrawn to a place where it is not likely to be damaged by blasted rock or should be suitably protected. The heavy blasting in mechanised quarried should take place during rest interval of workers.
The electrical circuit of detonators is tested for continuity by a circuit tester which is essentially a galvanometer. When everything is in order, the cable from detonator leads is connected to an explorer and the explosives are blasted from a protected place.
Pump-Truck Arrangement of Maharashtra Explosives Ltd., for SMS System:
In opencast mining where blasting is on a large scale, the SMS system is adopted as a standard practice, involving use of specially designed pump-trucks for transport to the blasting site ingredients required for SMS System.
The Site-Mixed Slurry System (SMS), basically comprises a mother support plant where an intermediate non-explosive slurry is, initially, prepared for SMS application. This intermediate slurry subsequently, is transferred to a 10-te capacity stainless steel tank mounted on a 18-te capacity, three-axled Ashok Leyland’ chassis, suitable for off-highway applications.
On rear side of the pump-truck, two small stainless steel tanks of 150-litres capacity each, are mounted which store cross-linker and sensitiser solutions. Each of these tanks is connected to a stainless steel screw pump powered by a hydraulic motor.
The intermediate slurry, cross-linter and sensitiser are simultaneously fed onto a mixing hopper at a pre-calibrated rate, wherefrom it is pumped into blast-holes with the help of progressive cavity stainless steel pump. The pump delivery is connected to a 15 m ling, static-charge resistant chemical hose of 50 mm diameter. All the control, check, relief and regulating valves for the hydraulic fluid are mounted at the rear end of the chassis in a control console.
For safety consideration, all storage vessels are made of stainless steel, the discharge hose is antistatic. The exhaust pipe is fitted with a flame-arrestor and a fire screen is provided separating the main body of the pump truck and the cabin.
Computer Blasting Model:
The latest technique in blast design is to use computer simulations which take into account rock properties, blast geometry and explosive characteristics. One such computer model is SABREX, marketed by ICI (India) Ltd. SABREX stands for Scientific Approach to Breaking Rock with Explosives.
The essential requirement for running Sabrex simulations is:
1. Five rock properties; Density, Young’s Modulus, Poisson’s ratio, compressive strength and tensile strength.
2. Explosive characteristics like shock energy, gas energy VOD, detonation pressure, density, etc. A companion programme called CPeX – Commercial Performance of Explosives – calculates these properties for a given composition and density.
3. Blast geometry in terms of hole diameter, bench height, burden, spacing, charge, stemming, delay pattern, etc.
Given these inputs, SABREX predicts fragment size analysis, throw, muck pile profile, damage envelope, fly rock and cost of drilling blasting. The results are displayed in colour graphics and tabulations. The blasting engineer can experiment with adjustments to many factors in a computer safety and achieve results of practical benefit.
1. SABREX accommodates variations of input to all elements of the blast design.
2. SABREX crack pattern is a view of the crack extending from each borehole yielding information on fragmentation, delay periods and back-break.
3. SABREX gives a picture of back break damage.
4. SABREX calculates fragmentation in terms of the percentage of passing size.
5. SABREX calculates ‘heave’ and builds muck pile profile to assist in subsequent digging operations.
Secondary Blasting:
Secondary blasting is carried out in two ways:
1. Pop Shooting:
A hole is drilled by jackhammer for charging with explosive and blasting the boulder. Normally a depth of 0.3 to 0.6 m is sufficient for most of the boulder sizes. The explosive, widely used, is special gelatine in conjunction with safety fuse or detonators.
2. Plaster Shooting:
A charge of explosive consisting of either a single primed cartridge or a few cartridges is laid on the surface of the boulder. It is then covered with a shovelful of plastic clay which is pressed into position by hand. It is advantageous to wet the surface of the stone before plastering and the clay should be well pressed down for good contact with stone round the explosive.
Special gelatine, or Ajax G, can be used for the blasting. I.E.L. has developed an explosive known as “Plaster Gelatine” for this purpose. It is a high velocity, high strength, gelatine type explosive suitable for such work. Plaster gelatine is not in the regular products of I.E.L. but can be available on request I.O.L. also markets a LOX cartridge which is used for contact blasting.
In plaster shooting the explosive charge used is about four times that required for non-shooting.
The object of secondary blasting is to break oversize boulders produced during the first or primary blasting to a size suitable for loading.
Blasting Gallery Technique:
This is a new technique for depillaring of thick coal seams.
Blasting gallery technique was adopted at GDK 10 incline of Singareni Collieries Co. Ltd. with collaboration of M/s Charbonnage de France. The method involves drivage of level galleries, and depillaring by drilling holes in a fan pattern around the gallery. The full thickness of the seam is extracted by retreating along the galleries, while blasting and loading out of successive slabs.
Due to the use of longer, the method requires special explosives and accessories. The explosives charge (of upto 3 kg) is distributed in the hole with the help of spacers, and non- incentive detonating fuse is used to ensure reliable of the explosives column.
ICI has developed special explosive – UNIRING – and a non-incendive detonating fuse – RINGCORD – which have passed special tests and have been approved for use with this method.
i. Coyote Blasting:
In this system a large quantity of explosive charge, nearly a few hundred tonnes, is packed in the large chambers inside the underground mine. These chambers are made by driving tunnels drifts or raises and are called coyote chambers.
After packing explosives in the coyote chambers, the connected tunnels/drifts/raises are backfilled tightly with the muck excavated when forming the chambers and the charge is normally blasted with the help of detonating cord.
ii. Cast Blasting:
Throw blasting or cast blasting is described as the controlled placement of overburden by drilling and blasting to achieve the optimum overburden removal cost. The technique is suitable for use with large or small draglines and can be adapted for shovel and dragline operations in conjunction with tract dozers to place upto 40% of the overburden in final position.