blast design for drifting and tunneling with wedge and burn cut

Blast design for construction of Tunnels or Drifts with Wedge or Burn Cut (Parallel Holes)

Author: Partha Das Sharma, (B.Tech-Hons. In Mining Engg.) Email: [email protected], Website: http://miningandblasting.wordpress.com/

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Blast Design for construction of Tunnels or Drifts with Wedge or Burn cut (Parallel Holes)

*** Partha Das Sharma (B.Tech-Hons. In Mining Engg.)

E.mail - [email protected] and [email protected] . Websites: http://miningandblasting.wordpress.com/ and

http://saferenvironment.wordpress.com. 1. Introduction – Tunnels, drifts, caverns and various underground excavations have been made in the earth for centuries by mankind for a myriad of uses. Excavations such as Ajanta, Ellora etc., are the well known examples of prehistoric era. In modern days, underground hard rock excavation and Tunneling is highly specialized type of work requiring very special type of explosives and drilling & blasting techniques. The size and scope of some of the projects can be tremendous and important for nation building such as construction of Subway systems for underground highways, Railways & mass transport, hydroelectric installations, water diversion system etc. As these projects are very special in nature, careful consideration is required in selection of explosives, selection of drills and their dia., designing of each round of blast to meet the specific site conditions, project constraints and economy. The blasting engineers ultimate challenge is to provide desired fragmentation with sufficient advance / progress and at the same time keeping ground vibration level within limit as per prevailing law. The controllable factors are explosive type, charge weight & geometry, distribution within a given rock mass. The uncontrollable factors are geologic condition of rock mass, environment condition and existing law of the land. The blasting engineers’ main challenge is to encounter those uncontrollable factors effectively by varying and proper utilizing controllable factors with desired economy. The development of tunnel construction techniques in recent years has been impressive. The prime objective in Tunnel blasting is to obtain maximum advance/pull per round of blast and to keep cost within reasonable limit. Therefore, very cautiously the type of explosives, drilling pattern (Wedge Cut or Parallel Holes with Dia. of empty holes), spacing & burden, number of holes to be drilled per round, Delay sequence etc., are to be selected. Cycle time is to be kept minimum as far as possible. 2. Cycle of operation and Parameter of Blast design - Drilling and blasting technique, owing to its capability to meet wide variations in geology, is predominantly used to excavate majority of the tunnels. Therefore, proper blast design in tunnels is the key to enhance blasting efficiency and reduce damage to the tunnel wall, vibration and noise level. Important tunnel blast design variables are burden, spacing, drilling dia, empty hole dia, face advance and tunnel area. Cycle of operation include Drilling, Charging, Blasting, Ventilation, Scaling, Support work, Grouting, Loading and Transport, and Setting out for the next blast. The factors on which a great deal of tunneling operations depends are:

1. Type of explosives used for tunneling blasting operations. 2. Blast design and selection of dia. & location of holes in compatible with the geology of

strata, designed area of opening, Environment, existing laws etc.

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The parameter of blast design for tunneling mainly depend on the factors like Geology of rock condition of the area and the Technique of the blasting adopted which again depend on Area of cross section of tunnel, Drilling & type of cuts, Delay sequence etc. 3. Geology, Rock condition & Site Characterization - Tunnel is one of the most hazardous projects in Engineering and construction. It is also one of the most expansive. Therefore, extensive planning, surveying, testing of ground samples etc., are done in pre-excavation stage of the project. There exist a number of uncertainties and unknowns about the site/ area and these become the major challenges to the designers of tunnel. A list of some of the most concern areas is: * Tackling uncertainties when dealing with any underground project * Geology of the area determines the feasibility and the cost of the project * Engineering properties of rock may change drastically with wide range of conditions, rate and direction of loading etc. * Ground water is the most difficult parameter to predict and most troublesome during construction * The most common method to determine underground rock condition is by core drilling, which only recover about 0.0005 % of the excavated volume of the tunnel with most exhaustive survey works, which leaves a great deal of room for uncertainties. Probably, the first activity in actual design phase is characterization of the site of the excavation and rock mass into which tunnel is to be driven. These include a) Topography of the area & the climate; b) Location of underground structure with respect to the ground surface and rock formation boundaries; c) Structural stability of rock body, presence of faults and stress concentrations; d) Hydrology of the area i.e., permeability of the ground and groundwater flow rate; e) Rock type, mass, their genesis and their homogeneity; f) Degree of weathering & weatherability of rock; g) Geologic discontinuities and other defects; h) Deformability characteristic under short and long term loading; i) In-situ stress and hydraulic / dynamic loads; j) Geometric and mechanical properties of systematic and extensive discontinuities. The most important is the rock must fulfill its ability to remain stable, when excavated. Therefore, it is important to consider the joints and cracks in a rock as these discontinuities can serve as a point of failure in a rock mass under stress. Tunnel engineers generally classify rocks on the basis of resistance to deformation (Strength), amount of weathering and general resistance to weathering. It has been seen that, Igneous and Metamorphic rocks are more resistant to deformation and weathering than Sedimentary rocks. 4. Sequence of excavation of various underground structures - Sequence of excavation of various underground structures depending on area of cross section of the excavation is given in Fig –1. Partial-face-blasting is sometimes more practical or may be required by ground condition or equipment limitations. Thus, tunnel-driving methods can be divided as: a. Full-face tunneling method, b. Top heading and bench method, c. Benching with horizontal blast hole, d. Benching with down holes.





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Fig-1

For tunnels generally smaller than 60 – 70 m² cross-sections, full-face excavation gives maximum economy and efficiency. Full-face excavation is applicable for tough rock having little jointing. Blast holes are drilled into tunnel face either at right angles to the face (Burn Cut) or at an angle to the face (Wedge Cut). For good roof condition and with availability of better support system tunnel having larger cross-section than 60 – 70 m² also can be effectively excavated by full-face method. For medium tunnel more than 60 – 70 m² area of cross section, normally top heading and benching sequence are adopted. In this method, driving an upper heading across the full width, one third to half of the final tunnel section. The lower section is removed later by benching. Top heading is generally driven throughout the full length of the tunnel before benching begins. In some operations the bench blasting is carried out simultaneously, but at another location within the tunnel. For excavation of very large underground cavern etc., more number of stages is employed. There are many other variations such as, a centre crown drift, followed by two crown side drifts, then bench in one, two or three stages. 5. Technique of the blasting - Tunneling in rocks is currently performed mainly by blasting, as this method only is capable of providing sufficiently high effectiveness and economics in the construction of tunnel in tough rocks. Tunneling by ‘tunnel borers’ is considered to be less effective especially as regards the construction of tunnels of large cross sectional areas.





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The blasts in tunnels and drifts are characterized by lack of adequate free surfaces towards which breakage can occur effectively. Unlike bench blasting, tunnel blasting has only one free face and holes are drilled normal to the free face surface. In such a situation, the explosives charge will blow out a narrow funnel- shaped crater. But if the hole is drilled at a certain angle to the free face, the result will be better, as the major part of the gasses will break out the rock in the direction of free face. Alternatively, if large diameter dummy holes parallel to the blast holes are drilled, the breakage performance is better as the large diameter dummy holes provide additional free face. Thus, the principle behind tunnel blasting is to create an opening by means of a cut ( a set of holes that provide initial free face) and then stoping to enlarge the opening. The cut, usually, has a surface area of 1 m² − 2 m², although with large drilling dia holes it can reach up to 4 m². The different zones in tunnel blasting are shown in Fig-2.

Fig-2

The initial opening/cut created either by angled holes or by holes drilled parallel to large diameter dummy holes are widen subsequently by the holes fired after cut holes using proper delays. In other words, the main difference between tunnel blasting and bench blasting is that tunnel blasting is done towards one free surface, while bench blasting is done towards two or more free surfaces. The rock is thus more constricted in the case of tunneling, and a second free face has to be created towards which the rock can break and be thrown away from the surface. This second face is produced by a cut in the tunnel face, which can be a parallel hole cut, a V-cut, or a fan-cut. After the cut opening is made, the stoping towards the cut begins. Stoping can geometrically be compared to bench blasting although it requires powder factors that are four to ten times higher. Such a high explosive consumption is mainly due to drilling error, the demand made by swelling, the absence of hole inclination, the lack of cooperation between adjacent charges and also blasting against gravity in case of lifter holes.





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The final shape of the cross section is given by trimmers or contour holes with closer spacing and comparatively smaller charge. Contour holes are spaced closely (0.2m to 0.4m apart) and directed outwards to make room for the drill in collaring and advance. The position of the cut has influence on rock projection, fragmentation and also on the number of blast holes. Of the three positions, namely, corner, lower centre and upper centre, the latter is usually chosen as it avoids the free fall of the material, the profile of the broken rock is more extended, less compact and better fragmented. 5.1. Most Common Cut Types for Tunneling are: * Wedge Cuts are largely banned in numerous operations as they eject rock violently, the rock is thrown a considerable distance resulting in services being destroyed e.g. ventilation, power, air and communications. * Frequently preferred cut type is the Burn Cut - Numerous variations exist and this is largely a ‘personal preference’ of the Blasting Engineer or Shotfirer. In other words, in tunnel blasting, selection of a cut plays a crucial role in achieving an optimized blasting. It can be done based upon rock characteristics, the area of tunnel and cut specifications and their applications. There are two large groups of cuts: parallel hole cuts and angled hole cuts. The first group is mostly used in operations with mechanized drilling, whereas those of the second have fallen in disuse due to the difficulty in drilling and being time-consuming and costly. The Cut is Critical for a successful blast – (a) Design the Cut correctly, (b) Position the Cut correctly on the face, (c) Drill the Cut correctly, (d) Load the Cut correctly, (e) Stem the Cut correctly, and (f) The Cut will ‘ pull’ and result in a good blast. If the Cut does not ‘pull’ the blast advance will be poor. A substantial part of the blast will ‘freeze’. Correcting this failure costs time and money. The Cut can be placed centrally or right or left of tunnel centre. The Cut should be placed about 1.5 to 1.8 m above the ‘Lifters’. A lower Cut position and early firing Lifter blastholes results in improved ‘floors’ and maximizes the effect of ‘gravity’ on the remaining blast portions. 6. Wedge Cut or V- cut - Horizontal cut holes are driven in inclined at an angle 45 to 60 degree to the face towards the centre. Maximum concentration of charge at the apex of the cut holes, which are fired first to create a free face for the rest of the shot, which are fired next with the help of delays. If cross-section of tunnel allows, for deep pulls, burden should be reduced by giving another shallow wedge cut known as 'Baby Cut' (Double V-Cut). These relieving holes (Baby Cut) should be fired prior to the main wedge cut pattern (Fig –3).





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Fig- 3

7. Parallel Hole Cut (Burn Cut) - The most commonly used cut in tunneling today is the large parallel hole cut, or Burn cut. Traditionally, the burn cut was drilled where the empty and loaded holes were all the same diameter. It was found that using empty hole of a large diameter than the loaded holes, provided additional relief in the pattern and reduced the amount of drill holes needed. The large empty holes also allowed for additional advance per round. A variety of names resulted from hybrid of the burn cut which used larger empty holes. In burn cut all holes are drilled at a right angle to the face and parallel to the tunnel direction. All holes in the large hole cut are drilled parallel to each other, and the blasting is carried out towards one or more empty large drill holes, which act as an opening. It is most important that Burn Cut holes be drilled parallel as possible and at the design distance from each other. Deviations can result poor blasting of the cut. The burn cut should be drilled at least 150 mm longer than the blast holes in the remainder of the round. The breakage is against the opening or void formed by the unloaded holes of diameter 76 to 150 mm. This opening is gradually opened up by successive detonation of the adjacent loaded holes and the pulverized rock is blown out of the cut. The cut may be placed at any location on the tunnel face, but its location influences the throw, the explosives consumption, and the number of holes needed in the round. If the cut is placed close to a sidewall, there is a probability of better exploitation of the drilling pattern, with less holes in the round. Furthermore, the cut may be placed alternately on the right or left side, in relatively undisturbed rock. To obtain good forward movement, and centring of the muckpile, the cut may be placed approximately in the middle of the cross section, and quite low down. This position will give less throw, and less explosives consumption, because of more stoping downwards. A high position of the cut gives an extended and easily loaded muckpile, with higher explosives consumption and more drilling, due to upwards stoping. The large empty hole cut comprises one or more uncharged large diameter holes, which are surrounded by small diameter blast holes with small burdens to the large holes. The blast holes are drilled in squares around the opening. The number of squares in the cut is limited by the fact





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that the burden in the last square must not exceed the burden of the stoping holes for a given charge concentration in the hole. The cut holes occupy an area of approximately 2 sq m. Small tunnel faces may need only cut holes and contour holes. When designing the cut, the following parameters are of importance for a good result: diameter of the large empty hole; burden; and charge concentration. In addition, the drilling precision is of the utmost importance, especially for the blast holes closest to the large holes. The slightest deviation can cause the blast hole to meet the large hole, or the burden to become excessively big. Too big a burden will only cause breakage or plastic deformation in the cut, resulting in lower advance. Design of Burn Cut holes - The overriding principle of all burn cut designs is as follows: Burden on loaded holes are selected so the volume of the rock broken by any hole cannot be greater than what would occupy the void space created by either the burn hole or subsequent holes firing. In this calculation one must also consider the fact that when rock web breaks between holes, it will occupy a larger space. In other words, the swell factor of the blasted rock must be considered. One of the parameters for good advance of the blasted round is the diameter of the large empty hole. The larger the diameter, the deeper the round may be drilled, and a greater advance can be expected. One of the most common causes of short advance is too small an empty hole in relation to the hole depth. An advance of approximately 95% can be expected for a hole depth of 4 m, and one empty hole with 102 mm diameter. If several empty holes are used, a fictitious diameter has to be calculated. The fictitious diameter of the opening may be calculated in accordance with the formula D = d √n, where D = fictitious empty large hole diameter; d = diameter of empty large holes; n = number of holes.

Fig - 4

In order to calculate the burden in the first square, the diameter of the large hole is used in the case of one large hole, and the fictitious diameter in the case of several large holes. The distance between the blasthole and the large empty hole should not be greater than 1.5 times the diameter of the larger hole for the opening to be clean blasted. If the distance is longer, there is merely breakage, and when the distance is shorter, there is a great risk that the blasthole and empty hole will meet.





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Fig - 5

So the position of the blastholes in the 1st square is expressed as: a =1.5d, where a = C- C distance between the large hole and the blasthole, d = diameter of the large hole. In the case of several large holes, the relation is expressed as: a =1.5D Where a = C- C distance between the centre point of the large holes and the blasthole, D = fictitious diameter. Therefore, side of the 1st square w1 = a√2 (Fig – 5). Position of blastholes in the 2nd square of the cut located at a distance of B1 from one of the sides of the 1st square, in such a way that B1 = w1 and C-C distance between the centre point of the large holes and the blasthole in the 2nd square is 1.5w1. Therefore, side of the 2nd square w2 = 1.5w1√2. Similarly, blastholes in the 3rd square of the cut located at a distance of B2 from side of 2nd square, in such a way that B2 = w2 and C-C distance between the centre point of the large holes and the blasthole 3rd square is 1.5w2. Therefore, side of the 3rd square w3 = 1.5w2√2. Similar calculation be followed for 4th square as well. The holes closest to the empty holes must be charged carefully. Too low a charge concentration in the hole may not break the rock, while too high a charge concentrate of ion may throw the rock against the opposite wall of the large hole with such high a velocity that the broken rock will be recompacted there, and not blown out through the large hole. Full advance is then not obtained. Generally, in average blastability rock, for 34 to 37 mm dia. blastholes in the 1st square are charged at a charge concentration of 0.5kg/m; for 2nd square blastholes charge concentration of 0.75kg/m; for 3rd square blastholes charge concentration of 1.15kg/m and for 4th square blastholes charge concentration of 1.25kg/m should be used.





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Fig - 6

It is a good practice to incorporate one or more large diameter uncharged relief holes together with smaller diameter charged blast holes in Burn Cut. Larger the diameter and more the number of uncharged relief holes in the cut, less serious is the consequence of drilling deviations, less chances of freezing and better advance per round of blast. Some of the burn cut patterns used in the field with more number of large diameter uncharged relief holes are given in the figures (Fig – 6, Fig – 7 and Fig - 8). One of the most common causes of short advance is too small an empty hole in relation to the hole depth. An advance of approximate 90% can be expected for a hole depth of 4m (45mm dia.) and one empty hole of 102 to 120 mm diameter.





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Fig - 7

The contour of the tunnel is divided into floor holes, wall holes and roof holes. The burden and spacing for the floor holes are the same as for the stoping holes. However, the floor holes are more heavily charged than the stoping holes, to compensate for gravity and for the weight of the rock masses from the rest of the round, which lie over them at the instant of detonation. For the wall and roof holes, two variants of contour blasting are used: normal profile blasting; and smooth blasting. With normal profile blasting, no particular consideration is given to the appearance and condition of the blasted contour. The same explosives as in the rest of the round are utilized, but with a lesser charge concentration and the contour holes are widely spaced. The contour of the tunnel becomes rough, irregular and cracked. The smooth blasting technique has been developed to obtain a smoother and stronger tunnel profile. Smooth blasting, where the contour holes are drilled close to each other and weaker explosives are used, produces tunnels with a regular profile, requiring substantially less reinforcement than if normal profile blasting is used.





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Fig - 8

The firing pattern must be designed so that each hole has free breakage. The angle of breakage is smallest in the cut area, where it is around 50 degrees. In the stoping area, the firing pattern should be designed so that the angle of breakage does not fall below 90 degrees. It is important in tunnel blasting to have sufficient time delay between the holes. In the cut area, the delay between the holes must be long enough to allow time for breakage, and throw of rock through the narrow empty hole, which takes place at a velocity of 40 to 60 m/sec. A cut drilled to 4 m depth would thus require a delay time of 60 to 100 milliseconds to be clean blasted. Normally, delay times of 75 to 100 milliseconds are used in the cut (Fig – 9).

Fig - 9

In the first two squares of the cut, only one detonator of each delay should be used. In the following two squares, two detonators of each delay may be used. In the stoping area, the delay time must be long enough to allow movement of the rock, to generate space for expansion of the





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adjacent rock to be loosened, say, 100 to 500 milliseconds. For the contour holes, the scatter in delay between the holes should be as small as possible, to obtain a good, smooth blasting effect. Therefore, the roof should be blasted with the same interval number, normally the second highest of the series, and here we can benefit from electronic detonators, as their scatter is practically nil. The walls are also blasted with the same period number, but with one delay lower than that of the roof. Detonators for tunnelling can be electric or non-electric. Electric detonators are manufactured as MS (millisecond) and HS (half-second) delays, and non-electric detonators as deci-second and halfsecond delays. Delay Sequence and use of NONEL: As mentioned earlier, blasting condition of tunnel is more difficult than bench blasting. The selection of delays and its timings should be such that, a free face is effectively created by moving out the broken rock mass so much that the rock volume after swell from subsequent blasts must be accommodated. To understand above the fracturing/ breaking time and time required to spreading out cracks in rocks to be studied. The time required to spreading out cracks and formation of fracturing/ breaking depend on the rock condition. The speed of developing cracks/ fractures is faster for hard and brittle rock than softer rocks. This rock breaking speed varies from 1 to 3 ms time per metre. Whereas, ejection speed of rock after blast may be about 20 to 30 metre per second, i.e., 2 to 3 cm/ms. Thus, for a 4 metre long hole broken rock takes about 300 to 400 ms for complete removal from the cut. For this reason, long period delays/ half-second delays are preferable for tunnel blasting. Some time, combination of short and long delays are used. The short delays mostly used in cut holes, as it requires shattering effect initially for rock breaking to create free face for easer holes.

Discussion on Design of Burn Cut (Parallel Hole Cuts) - The blasts in tunnels and drifts are much more complex than bench blasting owing to the fact that the only free surface is the tunnel heading. The powder factors are elevated and the charges are highly confined. On the other hand, burdens are small, which require sufficiently insensitive explosive to avoid sympathetic detonation and at the same time have enough detonation velocity (above 3000m/s) to prevent channel effect in the cartridge explosive placed in large dia blast holes. This phenomenon consists of the explosion gases pushing the air that exists between the column charge and the wall of the blasthole, compressing the cartridges in front of the shock wave, destroying the hot spots or excessively increasing the density of the explosive. Drilling has become more mechanized in the last decade with the introduction of fully automatic and computer assisted hydraulic jumbos with multiple booms. Thus, there is a general shift towards parallel hole cuts as they are easier to drill, do not require a change in the feed angle and the advances are not as conditioned by the width of the tunnels, as happens with angled cuts.





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Fig - 10

Some of the problems that can arise in blasting with parallel blasthole cuts are sympathetic detonation and dynamic pressure desensitization. The first phenomenon can appear in adjacent holes to the detonating hole when the explosive used has a high degree of sensitivity, such as, all those with nitroglycerine in their composition. On the other hand, the dynamic pressure desensitization takes place in many explosives and, especially, in ANFO. This is due to the compression of explosive charge ahead of the shock wave and resultant increase in the density of the adjacent charge above the critical density. Desensitization problem can be attenuated by correctly designing the initiation sequence, sufficiently delaying the successive detonation of each blasthole so that the shock wave from the last shot disappear, allowing the explosive to recuperate its normal density and degree of sensitivity. In order to minimise these problems, the parallel hole cut may be carried out by placing three relief holes in such a manner that they act as a shield between the charged holes. It has been observed that fine-grained rock is more susceptible to cut failure than coarse grained, due to the larger volume of relief opening that is needed for the expulsion of the material. Controlled Blasting - The ideal blast result in a minimum of damage to the rock that remains and a minimum of over break. This is achieved by controlled blasting, which has many advantages:

a) Less rock damage means greater stability and less ground support required. b) The tunneling operation is safer as less scaling is required. c) Less over break makes a smoother hydraulic surface for an unlined tunnel. d) For a lined tunnel, less concrete required to fill excess void created by over break.





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Controlled blasting involves a closer spacing of Contour holes or perimeter holes or Trim holes, which are loaded lighter than the Buffer and production holes. Generally, as a thumb rule 10 to 12 times hole diameter in medium to tough rock and 5 to 6 times hole diameter in poor, fragmented rock are kept as spacing. In fact, burden and spacing of Buffer and Contour row should be about 75% and 40% that of production rows respectively. Since, controlled usually require more blast holes, it takes longer to execute and uses more drill steel. For these reasons, contractors are often reluctant to incorporate principles of controlled blasting. Loading and charging of contour holes are done with explosives of low VOD packed in small diameter cartridges in relation to drill diameter used. Unlike production drill hole blast where higher charge concentration is required, contour drill holes require low charge concentration and explosives should be lightly distributed all along the length of the bore hole. Some time use of high grammage Detonating Fuse (about 40 gm/m core wt. To 60 gm/m core wt.) for contour blasting can give effective result in tunneling and civil construction works. This result in an air cushion effect, which prevent over break and reduces in-situ rock damage. More than just design of proper perimeter blast, other aspects of controlled blasting also to be looked at. Blast damage may occur long before the trim holes are detonated. Controlled blasting requires careful design and selection of all aspects of hole diameter, hole charge, hole spacing & burden, delays as well as careful execution of work. One of the keys to successful controlled blasting is precise drilling – deviation of holes from their design location leads to alter spacing & burden, causing blast damage and irregular surfaces. Other parameter to be looked at is charging of holes of Buffer row. In the buffer row the charge factor should not be more than 0.6 Kg/m³ to terminate the back break along the line of buffer row. At the same time, charge factor in the perimeter row should not be more than 0.4 Kg/m³ and it should be well distributed throughout the blast holes. Hollow bamboo spacers of 150 mm long may also be used to achieve better distribution of smaller cartridges in perimeter holes. A gradual reduction of charge factor from Cut holes at the center towards the perimeter of the tunnel will have efficient control on over break. Fig-8 shows charging of contour holes, buffer and production holes.

Fig-11





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Observation of the blasted surfaces after the blast can give better clues to the accuracy of drilling and blasting and effectiveness of the technique. A measure of success is the half-cast factor, is the ratio of half-casts of the blast holes visible on the blasted surface to the total length of trim holes. Depending on the rock quality a half-cast factor of 50 to 80 percent can be said to be quite satisfactory. There are other means to verify the extent of rock damage because of blast behind the wall. This may be done using Seismic refraction techniques and borescope or permeability measurements in cored bore-holes. The extent of disturbed zone may vary from as little as 0.1 – 0.2 m with excellent controlled blasting to more than 2 m with un-controlled blasting. ------------------------------------------------------------------------------------------------------------ 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, Mah. Explosives etc., before joining the present organization, Solar Group of Explosives Industries at Nagpur (India), few years ago. 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. 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], ------------------------------------------------------------------------------------------------------------------- Disclaimer: Views expressed in the article are solely of the author’s own and do not necessarily belong to any of the Company.

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blast design for drifting and tunneling with wedge and burn cut

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