# Chapter 35: Underground Blasting ## GENERAL DISCUSSION Underground excavations are created for one of two purposes. Either the rock itself has a value and a mine is developed if it can be processed at a profit, or the rock must be removed to create infrastructure space for some other purpose—in which case, the excavation is called construction. The drilling and blasting practices used to create the passageway in a mine or construction project are the same. The rock excavated to create these passageways generally has no value and is part of the mine development or construction project. The exception would be a room and pillar mine whose passageways are created as a result of the systematic mining of the rooms. As each room is mined, it becomes the access passageway to mine the next room. The terminology used to describe the direction of the passageway includes headings (horizontal direction), shaft sinking (downward direction) and raising (upward direction). Somewhat unique is the civil tunneling industry where, because of the large number of non-mining clients, headings tends to be called drifts, or when slightly inclined they are called ramps if heading down and declines if heading up as stoping and caving. The type (drill and blast) of the cut used depends on several factors, which include but are not limited to: (1) safe processes over the life of the mine, (2) the geology of both the ore and the host rock, and (3) cost effectiveness. When selecting the type of blast designate that will be utilized to create passageways or operate a mine, the most important basic information required is the structural geology of the area. The different rock characteristics (hardness, joints, seams, cavities) have a varying effect on how the rock reacts when it is blasted, and how stable the created opening will be after the blast. In sedimentary rocks, the bedding planes as illustrated in figures 35.1a and 35.1b are usually easy to identify, whereas in metamorphic or igneous rock, the type of jointing may vary considerably as illustrated in figures 35.2a and b. ![Figure 35.1a – Underground passageway intersecting thin horizontal bedded sedimentary rock (Tamrock, 1986).](images/716.png) ![Figure 35.1b – Underground passageway intersecting thin dipping bedded sedimentary rock (Tamrock, 1986).](images/717.png) ![Figure 35.2a – Underground passageway intersecting highly structured rock-large blocks (Tamrock, 1986).](images/717.png) ![Figure 35.2b – Underground passageway intersecting highly structured rock-small blocks (Tamrock, 1986).](images/718.png) --- ## Work Cycles The work cycle in underground mines and construction projects is a cyclical function in which different phases follow each other in repeating order. The three (3) primary activities are (1) drilling the boreholes; (2) blasting, (loading explosives, stemming, tying in, initiation, and ventilation) and (3) mucking, (scaling loose rock and excavating the blasted material). One complete cycle is called a "round." A round should produce a pre-designed advance of excavated length. In structurally competent rock, only these basic work phases may be required. However, the poorer the rock quality, the greater the focus must be directed to supporting the newly created opening as a result of blasting. Other support activities that may be required are installation of fresh air ventilation pipes and hoses, compressed air lines and placing rock support. The time required for each of these activities must be considered in an excavation schedule or plan. --- ## Explosives Underground blasting is generally carried out in a comparatively confined space that is often wet. Therefore, it is desirable to choose an explosive that has both good fume characteristics and water resistance. Some underground blasts can take many weeks to load. Although may not be obvious when the explosives loading process begins, water can slowly seep into the loaded boreholes and can be diverted to other holes as a result of the loading process. Both of these problems can cause non-water resistant explosive to become desensitized. The possible cost savings of using explosives that are not water resistant has to be weighed against the possible failure of any blast. For this reason, water resistant dynamites, emulsion and water gel explosives are preferred. However, when conditions are dry, ANFO primed with a cast booster, cartridge of dynamite, water gel or emulsion explosive is an excellent alternative. --- ## Initiation Systems Most blasting done underground utilizes either a conventional electric or nonelectric initiation system. However, the use of programmable electronic detonator systems has unique advantages for some applications, as has cap and fuse, which is still used in many areas of the world. The main consideration is the safe and reliable initiation of the explosive charge at the lowest cost. --- ## After Blast Fumes Underground air is subject to contamination by dust, smoke and fumes from various sources, such as internal combustion engines, rock handling equipment and blasting. In the United States, mining regulations are aimed at ensuring an adequate supply of clean air is supplied to work areas so as to reduce the health hazard of being exposed to these conditions. In most blasting operations, the required time lag between blasting and reentry to the work area is adequate for blasting fumes and suspended dust to be diluted or removed from the area by the ventilation system. Commercial explosives produce between 10 liters to 20 liters more carbon monoxide (CO) than nitrogen oxides (NO_x), but NO_x is 17 times more toxic than CO according to the threshold limit value (TLV) in table 2 of the American Conference of Governmental Industrial Hygienists. These components of post detonation gases represent roughly equal potential health hazard. Their behavior in an underground environment is different. Detonation of an explosive produces nitrogen oxide (NO) rather than nitrogen dioxide (NO₂), but in the presence of atmospheric oxygen, the NO oxidizes to NO₂. This gas is soluble in water, and if any moisture is present, the NO₂ concentration usually declines very quickly. Since CO is chemically stable and is only slightly soluble in water, it is more persistent. If personnel operate that area too quickly, the smell of NO₂ gives a warning that dangerous fumes are present. With enough time the NO₂ is likely to diminish through absorption in water, even with inadequate ventilation, but the insoluble CO, which is dependent on ventilation for dilution. Therefore, smell is not a reliable method of determining whether the air has dangerous levels of fume. Following the precautions in table 35.1 ensures that personnel are not exposed to unhealthy levels of fume. ### Precautions To Prevent Personnel Exposure To Unhealthy Levels Of Fume | Precaution | |------------| | Periodically measure the working place for fumes. | | Ensure good ventilation | | Observe appropriate reentry time periods. | | Use a water spray to suppress dust and fumes | | Choose an explosive with adequate water resistance | *Table 35.1 – Precautions to prevent personnel exposure to unhealthy levels of fume.* --- ## Safeguarding the Blast Area The most frequent cause of underground blasting accidents is improper blast area security. A system for clearing all personnel to a safe location and safeguarding of equipment prior to initiating a blast must be planned and followed. After initiating the blast, the area should not be reentered until the ground has stabilized and ventilation has rid the area of all smoke and noxious gases. --- ## OVERBREAK CONTROL The names given to the various techniques vary, but the term final wall blasting (controlled blasting) refers to all methods which have a purpose to reduce and distribute the explosive charge energy level in the final line of blast holes that defines the perimeter of the intended excavation so as to minimize cracking beyond the perimeter and thus improve stability of the rock that is to remain. The final wall blasting techniques discussed in chapter 36 are applicable to any type of rock. However, geologic structures, as planes of weakness, have a tendency to shear or fracture beyond the perimeter. If the rock is weak, possibly no blasting technique can create a smoother, more solid face than permitted by the inherent rock formation. The best results occur in homogeneous formations having a minimum of geologic structures and seams. The general strategies to control overbreak at the perimeter of an excavation are summarized in table 35.2. ### Blasting Strategies To Control Overbreak | Strategy | |----------| | Reduce charge weights in the final line of blast holes. | | Drill extra holes to compensate for the reduced charge weight/borehole and provide a preferential plane of weakness to delineate the final excavation perimeter | | Arrange the initiation of the boreholes defining the perimeter to obtain maximum relief and movement toward the interior of the excavation. | *Table 35.2 – Blasting strategies to control overbreak.* The two most widely used methods of overbreak control in underground blasting are (1) presplitting and (2) smooth blasting. These techniques are summarized in table 35.3. Smooth Blasting is the most widely used overbreak control method in shaft and tunnel excavations. Presplitting is utilized in drifting during bench and stope-hole excavations (See chapter 36). ### Most Commonly Used Techniques To Control Overbreak In Underground Blasting | Technique | Description | |-----------|-------------| | Presplit | Detonates the perimeter boreholes simultaneously and ahead of the inner boreholes. | | Smooth blasting | Reduces the energy level in the perimeter boreholes. Use an appropriate loading configuration.* | *Smooth blasting normally uses a burden dimension between 1.3 times and 1.5 times the spacing. *Table 35.3 – Most commonly used techniques to control overbreak in underground blasting.* Perimeter boreholes are normally initiated last in the delay sequence and rely on a reduced charge weight per borehole to control overbreak. When horizontal boreholes are loaded using thin technique, it is recommended to secure the explosive cartridges with a tamping plug or other device to prevent them from being dislodged from their respective boreholes by the action of earlier detonating charges of the primary blasting round. Table 35.4 lists recommendations for perimeter hole spacing and burden. These recommendations are for generalized conditions and should be adjusted as rock conditions change. ### Perimeter Borehole Burden and Spacing Dimensions vs. Powder Factor | Borehole diameter (millimeters) | Spacing (meters) | Burden (meters) | |--------------------------------|------------------|-----------------| | 32 | 0.40-0.50 | 0.50-0.65 | | 38 | 0.45-0.60 | 0.60-0.80 | | 45 | 0.55-0.70 | 0.70-0.90 | | 51 | 0.65-0.90 | 0.85-1.15 | *Table 35.4 – Perimeter borehole spacing and burden dimensions vs. powder factor.* Whichever technique is used, the factors summarized in table 35.5 must be considered. ### Factors Affecting Overbreak | Factor | Description | |--------|-------------| | Drilling accuracy | Toe spacing error should not exceed 50% of the planned spacing. Non parallel holes will result in loss of profile and impaired breakage at the perimeter. The undermining of sidewalls by poorly drilled, and heavily charged corner Rib or boreholes sometimes causes deterioration of the perimeter in tunneling. | | Subdrifting | Subdrifting is generally required to excavate to floor grade when bench blasting. If subdrifting is utilized at the perimeter of the bench along a pillar, the user centered charge below grade level actually resides in the stability zone. | | Coupling Ratio (See Chap. 11) | In order to prevent damage while providing positive breakage, the diameter of the charge should be 1/2 to 2/3 times the borehole diameter. Experience has shown that when the explosive diameter is less than half that of the borehole diameter, the coupling ratio is less important than the total charge weight in the hole. A smoother profile of the remaining rock face is obtained with uniform coherent charges. | | Stemming | Stemming also prevents the cartridges from being permanently ejected as a result of gas migration from inner boreholes. Stemming can sometimes produce surface results (in very weak rock). Here the confined explosive gases can be forced into preexisting fissures. As a general rule, effective stemming length of 10 times the borehole diameter is adequate. | | Guide holes | In extreme conditions or accord caverns, an improved finish can be obtained by drilling extra, uncharged holes along the desired perimeter line. These "guide holes," serve to direct and contain cracks growing from charged holes. A useful modification of guide holes is the practice of using only one primer for the 3 hole group that is illustrated in figure 35.3. | *Table 35.5 – Factors affecting overbreak.* --- ## TYPES OF CUTS The boreholes that create the opening cut are the most important part of any blast design. The remaining boreholes in the round cannot break effectively to their full depth unless the "cut" pulls to the bottom and the fragmented rock is ejected from the face. Therefore, the "cut" must create an empty space towards which the rest of the round is blasted. To help improve the "pull" of a cut, those boreholes should be drilled slightly deeper than the rest of the boreholes in the round. There are many types of opening cuts being used in underground blasting. For classification purposes, they are grouped into two categories: parallel hole cuts and angle cuts. --- ### Parallel Borehole Cuts In parallel borehole cuts, all boreholes are drilled parallel to each other (perpendicular to the face) around one or more holes not loaded with explosives. This cut is most commonly used in tunnels, shafts and raises. The boreholes are closely spaced around the empty relief hole(s), which provide a void space for the initial boreholes to sequentially detonate and gradually enlarge the opening, creating a larger space for the rest of the blast, the determining factors being geometrical rather than maximizing blasting technical values. The spacing and loading of cut holes is critical to achieving an advance of 90% to 95% of borehole depth. Generally, the center-to-center distance from a single empty hole to the first borehole of the delay sequence should be about 1.5 times the diameter of the empty hole. Thus, for an empty relief hole diameter of 100 millimeters (4 inches), the spacing of the nearest blast hole should be about 150 millimeters (6 inches). Also, if the spacing dimension of the loaded hole is greater than two times the empty hole diameter, plastic deformation of the rock occurs, the bottom of the cut does not clean out and the rest of the round fails to pull to the design depth. Figure 35.4 shows the potential success of blasting at varying distances from different size unloaded holes. ![Figure 35.4 Result when blasting toward an empty hole at different distances and dimensions of the empty hole (After Langfors and Kihlstrom). (Source: ISEE Blasters' Handbook™, 17th Ed. figure 29.9)](images/722.png) Note: if more than one empty hole is utilized, the spacing dimension to the first borehole can be increased and the probability for success is improved. ![Figure 35.5 – Loaded borehole vs. empty relief hole layout. (Source: ISEE Blasters Handbook™ 17th Edition, figures 25.7 and 25.9)](images/723.png) Figure 35.5 shows several commonly used relief hole configurations. Parallel hole cuts have greater flexibility than angle cuts by allowing changes in the depth of the round without changing the drilling pattern or direction of the holes. Thus, changes in the advance per round can easily be accommodated to adjust cycle times. Parallel hole cuts generally have a higher relative advance (percent of drilled depth) and result in better fragmentation and less fly rock than angle hole cuts. Parallel hole cuts are generally called "box cuts", although the term "starter cut" and "cylinder cut" are sometimes used. --- ### Angle Borehole Cuts Angle hole cuts are generally used only on faces with relatively large cross-sectional areas, such as in room and pillar mining applications where the heading is wide enough to allow the drills to be placed at angles to the face. ![Figure 35.6 – Angle hole drill setup-underground. (Courtesy: Austin Powder Company)](images/724.png) Therefore, the advance is somewhat dependent on the width of the opening. In addition to the restrictions caused by heading width, there is also the issue of accurate hole alignment and depth due to the many changes at the drill hole depths angle. Angle cut designs generally require fewer holes per round and less explosives per unit length of advance than parallel hole cuts. There are two basic types of angle cuts: (1) V-cut, and (2) fan cut. --- ### V-cut The V-cut is based on symmetrically drilled angled holes and is the most commonly used type of cut in room and pillar mining. The main limitation of the V-cut is the drilling of the "V" itself. Unless the internal angle of the "V" is greater than 60°, achieving consistent good results is difficult. Also, the amount of advance is limited to 86% of the V-cut borehole depth. ![Figure 35.7 – V-cut drill hole alignment, hole depth, hole spacing. (Adapted from ISEE Blasters' Handbook™, 17th Ed. figure 21.15)](images/725.png) The cut may have one or several "V's" drilled parallel to each other as in figure 35.7. The number of "V's" utilized in a design depends upon the structure or or stratification of the rock. In deeper rounds or in hard-to-break rock, the cut may consist of multiple "V's" and could also include a smaller "V" called the "baby-V." "Burster" holes are sometimes used to help break up the large mass of rock formed by the action of the "V," in deeper cuts as illustrated in figure 35.8. To produce the best results, all the boreholes in the V-cut should be fired simultaneously using a "zero" or a 25 millisecond delay. The time required for burden movement must also be given consideration so as to create adequate void space for the unit boreholes as the delay pattern sequence. Figure 35.9 illustrates and suggests timing for the V-cut round design. ![Figure 35.8 – V-cut with baby "V" and burster boreholes (plan view). (Source: ISEE Blasters' Handbook™, 17th Ed. figure 29.7)](images/725.png) Boreholes not drilled to the same vertical plane waste drill footage and explosives. As well, more time is required to scale a damaged face without any increase in advance. Also, the muck pile will normally be spread farther down the opening when using angled cuts. This could damage the ventilation system or other equipment from flyrock. ![Figure 35.9 – V-cut pattern with spacing dimensions and delay sequence (front view). (Courtesy: Austin Powder Company)](images/726.png) --- ### Fan Cut Fan cuts are usually applied to very soft rock formations, such as talc, gypsum, coal and salt with or without a "kerf" or where large mechanized drilling equipment cannot make the 90° turn to "square" the machine to the face for the first round in room and pillar mining. ![Figure 35.10 – First fan round turn with drill jumbo. (Tamrock, 1984)](images/726.png) To fully utilize a fan drill pattern, the drill pattern must be alternated from side to side to keep the heading "square" to the centerline of the passageway as illustrated in figures 35.11 and 35.12. ![Figure 35.11 – Succession of three fan rounds. (Source: ISEE Blasters' Handbook™, 17th Ed. figure 21.28b)](images/727.png) ![Figure 35.12 – Swing cut. (Source: ISEE Blasters Handbook™ 17th Edition, figure 21.28a)](images/727.png) --- ## DEVELOPMENT BLASTING One of the most important parameters in planning the excavation cycle is the length of the round. It has a noticeable affect on all the other work phases. For instance, a longer round means: (1) more drilled length of hole, (2) more explosives, (3) longer loading time, (4) larger volume of rock for mucking; and (5) more roof area to support between cycles. Long drilling depths produce longer advances per round and allow high capacities, in theory. In practice, however, the actual excavation capacity depends on the working arrangement and scheduling of other work activities. The practical advance (also called "pull") is usually limited to 90% to 95% of the drilled depth. The longer the round, the lower the percentage of advance will be achieved due to borehole deviation at the bottom of the boreholes which creates less predictable burden dimensions. The best blast designs are of little value unless the spacing and burden dimensions specified in the design are met. --- ### Drifting and Tunneling Methods The blast pattern designed to create a civil tunnels utilize the same principles as is used to develop smaller mining drifts—the difference is that drifts are made only as large as necessary for people and equipment to work and operate a mine, whereas a civil tunnel has different profile requirements and must last for a much longer time, and therefore, greater emphasis is placed on contour control. Tunnels vary in size and shape in accordance with the purpose for which they will be utilized. Those used for sewers and underground railways may be as small as 3 meters (10 feet) in diameter. Tunnel Boring Machines (TBMs), can excavate full-face tunnels up to about 14 meters (46 feet) in diameter and have replaced drilling and blasting as the most economical excavation method for tunnels over 1,000 meters (3,500 feet) in length. However, in shorter tunnel lengths, or in unusual geologic conditions, drill and blast continues to be utilized. For tunnels in rock, it can be shown that all cross sections with three or more rounded or curved corners require approximately the same number of boreholes. ### Most Commonly Used Tunnel Drilling and Blasting Methods | Method | |--------| | Full-face | | Top heading and bench | | Pilot tunnel | *Table 35.6 – Most commonly used tunnel drilling and blasting methods.* ### Factors Affecting Tunnel Drilling and Blasting | Factor | |--------| | Ground conditions | | Diameter and length of tunnel | | Available equipment | | Mucking method | | Adjacent structures of concern | | Physical location of project | *Table 35.7 – Factors affecting tunnel drilling and blasting.* --- ### Full-Face Method As the name implies, this method uses a round design to drill and blast the full cross-sectional area of a tunnel face in one blast. It is limited only by the capability of the drilling equipment. The procedure has always been utilized in small diameter tunnels, however, with the introduction of larger mobile drills, the full-face technique can be applied to practically all tunnels less than 9.1 meters (30 feet) in length unless the ground conditions require a different method. ![Figure 35.13 – Typical full-face round for a 4.8 meter (16 foot) wide horseshoe tunnel using two large diameter relief holes. (Source: ISEE Blasters' Handbook™, 17th Ed. figure 29.1)](images/728.png) --- ### Top Heading and Bench Method The top heading and bench is the standard method for tunnels greater than 9.1 meters (30 feet) in diameter. It is also used where capability of drilling equipment is limited. It consists of developing a heading at the top of the tunnel that takes in a portion of the finished height and its full width or a portion of it. After ground support is installed, the lower portion is then removed in one or more benches as illustrated in figures 35.14. ![Figure 35.14 – (Left) Top heading and benching method-vertical boreholes (front view) (Courtesy: Austin Powder). (Right) Top heading and benching method-vertical boreholes (profile view). (Source: ISEE Blasters' Handbook™, 17th Ed. figure 29.2)](images/728.png) Figure 35.14 illustrates the method using vertical boreholes and figure 35.15 shows the same method using horizontal boreholes for the benching operation. Whether vertical or horizontal boreholes are used in the bench, multiple rows are usually blasted in a delay sequence towards the open face of the bench. ![Figure 35.15 – (Left) Top heading and bench method – horizontal bench boreholes (front view) (ISEE Blasters Handbook™ 17th Edition). (Right) Top heading and bench method – horizontal bench borehole (profile view). (Courtesy: Austin Powder)](images/729.png) --- ### Pilot Tunnel Method A small pilot heading is driven portal-to-portal down the centerline of the proposed larger tunnel by conventional drilling and blasting techniques to provide the owner and the contractor a "look" at the geologic structure of the rock before enlarging the pilot tunnel. This method of tunneling is particularly useful where it is extremely difficult to acquire rock core borings and other sub-surface information. --- ### Tunnel Drilling Patterns The drill pattern designs and blasting process to create underground passageways (tunnels, shafts, raises) differ from other types of blasting because only one free surface is available—the surface from which the boreholes are drilled. Blasting is carried out by drilling and loading horizontal or nearly horizontal boreholes into the vertical tunnel face, then delaying them in rotation (generally from the middle of the heading outward to the perimeter). The basic idea in this type of confined blasting is to create an opening with a "cut" and then blast the rest of the rock towards that opening. ![Figure 35.16 – Pilot tunnel method. (Courtesy: Austin Powder Company)](images/729.png) ![Figure 35.17 – Cut in a tunnel round (Tamrock, 1984).](images/730.png) The "cut" is the most important part of any blast design. If it fails to function as planned, the remainder of the blast design also under performs. Delay timing in the cut for a burn round is critical to prevent the round from freezing. Delay timing for a 3 meter (10 foot) ranges from 10 milliseconds in hard competent rock to 50 milliseconds in soft rock per foot of round depth to allow the cut rock to clear the cut before the next boreholes are detonated, i.e. for a 10 foot deep round, use 100 milliseconds minimum between holes for hard rock and 500 milliseconds between boreholes for soft rock. After the cut is formed, cohesive formation boreholes are fired sequentially around the cut to enlarge the opening to the perimeter of the tunnel excavation. The purpose of these boreholes is to: (1) attain the same advance as the cut, (2) obtain satisfactory fragmentation, and (3) to throw the rock from the heading in a manner which allows quick and easy mucking without damaging utilities. The most complex overload control begins with the first row of boreholes in from the parameter as illustrated in the *Overbreak* section of this chapter. This "ring" of boreholes must create suitable burden conditions for the perimeter holes that are more closely spaced and lightly loaded in order to break to the design limits without overbreak and damage to the remaining rock. The last boreholes initiated in a round round are generally the lifter boreholes, located at the tunnel "floor." The charge concentration in these boreholes is generally greater because of the increased burden as a result of the required drilling of the boreholes below floor grade and the fact that at the time these holes are detonated in the delay sequence a large amount of the muck from the blasted rounds is in front of and on top of the lifters. Generally, drilling patterns are designed to produce consistent results within a predictable work cycle of time. From these borehole patterns, blasters can determine pertinent blast design parameters by image the area of the opening. As can be seen in figure 35.18 the smaller the cross section area of the passageway, the higher the powder factor is needed. Positioning the cut in the center of the round produces the best throw and centering of the muck pile. Placing the opening cut lower in the round can result in less throw and a tighter muck pile. Placing the cut higher in the face results in more spreading of the muckpile. ![Figure 35.18 – General relationship between powder factor and face area of a tunnel, shaft, or raise blast (Tamrock, 1984).](images/731.png) Figure 35.19 illustrates the relationship between the approximate number of boreholes and the cross section area of the tunnel. These curves are offered as average guidelines that will require adjustments to individual rock types and conditions. ![Figure 35.19 – Area per borehole for tunneling, shaft sinking and raise blasting. (Source: ISEE Blasters' Handbook™, 17th Edition)](images/731.png) In general, the largest diameter drill hole, consistent with available drill equipment and drilling rates, will offer the most economical drilling and blasting cost. This is especially true of large tunnels and rooms and less of a factor in smaller tunnels, where the decrease in the number of holes is offset by the need to obtain the appropriate tunnel shape without contributing to over break as shown in figure 35.20. ![Figure 35.20 – Number of boreholes per round for tunneling, shaft sinking, or raise blasting. (Tamrock, 1988)](images/732.png) --- ### Shaft Sinking and Raise Driving Methods An essential part of creating the proper access to underground mines and construction projects is the development of shafts and raises for a variety of purposes. Sinking a shaft does not require any existing underground access, whereas this is a vital requirement for development of a raise, which is driven from the bottom up. The blast design principles described for tunnel rounds may also be applied with some modifications for both shaft sinking and raising. As in tunnel blasting, the "cut" is critically important. ![Figure 35.21 – Shaft sinking, half bottom method (simultaneous sinking and benching rounds). (Source: ISEE Blasters Handbook™, 17th Edition)](images/732.png) --- ### Shaft Sinking Methods There are two basic half and blast shaft sinking methods: (1) Half-bottom benching method, and (2) Full bottom method. The benching method is particularly suitable for sinking square or rectangular shafts or where the cross-sectional area of the shaft is very large. The lower half serves as a sump as the upper half of the shaft cross section is drilled with a fan type drill pattern. The cycle is then repeated alternating down in a series of steps as illustrated in figure 35.21. In the full bottom method the entire cross section is drilled and blasted at one time. The cut is usually a burn cut design with angled holes (cone pattern) immediately surrounding the empty holes. For shallower shafts of depths less than 100 meters (325 feet), a very large diameter relief hole 200 millimeters to 450 millimeters (8 inches to 18 inches) is often drilled full depth prior to the start of sinking the shaft and serves as the "cut" for all of the blasting rounds. ![Figure 35.22 – Large shaft sinking drill pattern. (Courtesy: Austin Powder)](images/733.png) Figure 35.22 illustrates a very large relief hole that is backfilled with sand or pebble stemming so that it can be cleaned out to the depth of the next round when explosives are to be loaded into the blast holes as the shaft progresses downward. In some shafts, large pilot holes of 1 meter to 4 meters (3 feet to 13 feet) in diameter are drilled to intersect a tunnel or chamber at the bottom of the shaft (See figure 35.23). ![Figure 35.23 – Shaft sinking round with large pilot hole. (Courtesy: Austin Powder)](images/734.png) In addition to providing excellent relief, the large pilot hole provides ventilation, drainage, and quick means of mucking the blasted round. A popular large-shaft round design is the pyramid cut as illustrated in figure 35.24 and the V-cut shown in figure 35.25. ![Figure 35.24 – Shaft sinking round with pyramid cut. (Source: ISEE Blasters Handbook™ 17th Edition, figure 29.20)](images/735.png) ![Figure 35.25 – Shaft sinking round with V-cut. (Source: ISEE Blasters Handbook™ 17th Edition, figure 29.21)](images/736.png) As in tunneling, the "cut" should be deeper than of the rest of the boreholes of the design to maximize the potential of advancing the rock at the bottom of the "cut" and to provide a sump for pump placement. --- ### Raise Driving Raises are vertical or steeply inclined openings driven upward to connect levels, develop stopes, and provide access into upper areas. They vary greatly in size and inclination depending on their purpose. The methods used for drilling and blasting raises depend largely on the length of the raise and are: (1) Timbered raise building and Alimak method, and (2) Long hole and vertical crater retreat. --- ### Timbered Raise Building and Alimak Method Drilling blast holes spaced in successive blasting rounds use the same blast designs as driving small tunnels or drifts, except that the holes are drilled upward with hand held drills from a working platform. The pattern would normally not exceed 4 feet to shoot one, utilizing a smaller hole, small diameter booster and cartridges drilled using that drill size column loader. Since a closely spaced face is far safer than a cracked and partially pulled round, it is vitally important that the cut functions as designed. Mucking is achieved by gravity as each blast hole detonates in sequence. Miners return to the working floor after ventilation air replaces the smoke and fumes to clean loose rock from the timbers, and scale the face prior to drilling the next round. ![Figure 35.26 – Raise driving by Timber raise building and Alimak method (Tamrock, 1986).](images/736.png) --- ### Long-hole Method An alternative to creating raises in an upward direction with small holes is the practice of using larger diameter holes drilled and loaded with explosives from above. Long hole raises are safer and faster to develop since all the drilling and blasting is done from the level above. The rock falls to the bottom and is mucked separately from the blasting cycle. The major limitation with the long-hole method of raise blasting is the drill hole accuracy over long distances. After the holes are drilled, they should be mapped (See chapter 32) so that the delay sequence of the boreholes and charges can be changed as necessary at different elevations of the raise as a result of borehole deviation. Explosives are lowered into position from the upper level after the boreholes are blocked or plugged with wooden wedges, inflatable plugs, or other similar devices. Stemming is kept at a minimum because these boreholes are used over and over until the raise is completed. The two types of blasting that fall into this long-hole method category are the: (1) medium diameter borehole method and, (2) VCR (vertical crater retreat) drop raise method. ![Figure 35.27 – Long-hole raise method (Tamrock, 1986).](images/737.png) --- ### Medium Diameter Borehole Method The pilot opening is generally blasted using medium-diameter holes in rounds in depths of 2 meters to 4 meters (6 feet to 12 feet) that are enlarged to the design dimensions in subsequent blasts of 3 meters to 6 meters (10 feet to 25 foot) length. Figure 35.28 illustrates a 2.5 meter by 2.5 meter (8.2 feet by 8.2 feet) square raise design with 63 millimeter (2½ inch) holes delayed with Long Period delay detonators. The practical advance of this type of round is usually limited to the blast diameter of the pattern, as in tunnel and shaft rounds. Therefore, loading more than a 2.5 meter (8.2 feet) explosive charge length in this example would probably not produce more advance and could possibly "freeze" the cut. It is possible to advance multiple rounds as one blast by using stacked delay explosive decks, however if for any reason one (1) explosive charge does not function as designed, the entire raise could be lost. Therefore, the delay timing of the explosive decks and proper stemming between the decked charges is critical to the success of this blasting application. ![Figure 35.28 – Raise borehole layout and delay sequence. (Source: ISEE Blasters' Handbook™, 17th Edition)](images/738.png) --- ### Large Diameter Borehole Method Vertical crater retreat raises, commonly called VCR drop raises, utilize 152 millimeter to 165 millimeter (6 inch to 6½ inch) diameter boreholes drilled from the level above to the level below an open stope, or an existing tunnel. Instead of empty relief holes, the initial opening or cit is created by blasting a crater (See chapter 9) in the center of the drill pattern using spherical charge geometry (see VCR stoping) so that the other boreholes then sequence their breakage into that opening. The timing has been successful using both MS and LP delays. See figure 35.30 for a typical pattern. ![Figure 35.29 – VCR borehole performance (profile). (Source: ISEE Blasters' Handbook™, 17th Ed. figure 25.17)](images/738.png) ![Figure 35.30 – Typical VCR pattern and delay sequence. (Source: ISEE Blasters' Handbook™, 17th Ed. figure 25.18)](images/739.png) --- ## STOPING AND PRODUCTION BLASTING Underground blasts are commonly referred to as production blasts if they: (1) widen a tunnel, room, chamber or stope; (2) reduce or remove a pillar; (3) lower the floor in a tunnel, room or chamber; or (3) reduce caving. The basic concept for these types of blasting consists of one or more rows of holes drilled parallel to and blasted toward an area of open relief or a "slice" or "slab" round. These production blasts may be classified as either (1) short-hole or (2) long-hole blasts. --- ### Short-Hole Blasts Short-hole blasts are generally used in construction applications to enlarge a tunnel or chamber and in the mining of cut and fill, square set, shrinkage and top slice stopes. Variations of this type of blasting are referred to as overhand stoping and underhand stoping. The width and length of advance in a stope round depends on the structural integrity of the opened ground. --- ### Overhand Stoping Breasted are drilled with small diameter horizontal or upward angled holes and blasted down in a series ascending steps. The miners work on previously broken ore (shrinkage stopes) or back-fill material (cut and fill stopes) for the next drilling and blasting cycle. The overhand stoping blasting method is illustrated in figures 35.31. ![Figure 35.31 (Left) Overhand stoping method-vertical boreholes. (Right) Overhand stoping method-horizontal boreholes. (Source: ISEE Blasters' Handbook™, 17th Ed. figure 25.21)](images/740.png) ![Figure 35.32 – (Left) Overhand stoping-horizontal borehole delay pattern sequence (front view). (Right) Overhand stoping-horizontal boreholes (profile view) (Source: ISEE Blasters' Handbook™, 17th Ed. figure 25.22)](images/740.png) The process is repeated as the mining cycle follows the ore body upwards. Ore is drawn and/or backfill is added to maintain a working floor for miners. A typical cut-and-fill is shown in figure 35.32. --- ### Underhand Stoping Underhand stope blasts use downward vertical or angled holes to shoot the ore in descending steps or benches as illustrated in figure 35.33. ![Figure 35.33 – Underhand stoping method. (Source: ISEE Blasters' Handbook™, 17th Ed. figure 25.25)](images/740.png) --- ### Heading and Slab Rounds Many types of mineral and aggregate deposits are found in flat lying sedimentary and metamorphic geology, which allows the use of the room and pillar mining method. When a geologic formation is less than 3 meters (10 feet) thick the headings and crosscuts are usually mined the full height of the mineral deposit. The deposit is over 3 to 5 meters (10 ft thickness), it is accessed at the top of the geologic formation and then the floor is benched between the pillars after the top heading arrangement has advanced sufficiently. Figures 35.35 and 35.9 (See section *Angled Borehole Cuts*) illustrate a V-cut production round. ![Figure 35.34a – Production round with kerf (face view). (Source: ISEE Blasters' Handbook™, 17th Ed. figure 25.27)](images/741.png) ![Figure 35.34b – Production round with kerf (profile view). (Source: ISEE Blasters Handbook™, 17th Ed. figure 25.27)](images/741.png) Figures 35.34a and 35.34b illustrate a typical machine-cut kerf and borehole pattern used in potash. --- ### Long-Hole Blasts Long-hole blasts are used in the mining of an open stope (e.g. bench, sublevel, or VCR shrinkage methods) or undercut cave (sublevel or block) by a series rings and/or rows of medium to large diameter blast holes. --- ### Long-Hole Rings and Fans Boreholes are drilled radially from an access drift. All the rings are drilled parallel to a slot raise or open stope. The rings of blast holes may be full circle or partial depending upon the ore, the waste rock and the relief area. The number of rings blasted depends upon the purpose of the blast and ground stability required after blasting. Spacing between boreholes within a ring is greatest at the bottom end of the boreholes. This dimension will be dependent on the ore properties, fragmentation required, explosives used, and the diameter of the borehole. The key considerations in the placement of the boreholes and their bottom's ends is to assure that the powder factor and placement of the explosive will adequately fragment the rock between the ends of the boreholes. Figure 35.35 shows a typical ring of holes shooting to an open stope underneath and vertical slot along side the ring(s). ![Figure 35.35 – Ring blast borehole design (front view). (Tamrock, 1986)](images/742.png) ![Figure 35.36 – Vertical crater retreat stope. (Courtesy: Atlas Copco)](images/742.png) ![Figure 35.37 – Shows a typical loaded VCR borehole. (Source: ISEE Blasters' Handbook™, 17th Ed. figure 25.15)](images/743.png) --- ### Vertical Crater Retreat Blasts The VCR (Vertical Crater Retreat) mining method is a form of shrinkage stoping. This method is based on the use of a "spherical" charge rather than the cylindrical charge used in most blasting. A "spherical" charge exists when the length of the charge compared to its diameter does not exceed a ratio of 6:1. Before a VCR stope is blasted the top-sill, under-cut, and extraction points are all mined. The stope is completely drilled out from the top-sill with 152 millimeter to 165 millimeter (6 inch to 6½ inch) diameter boreholes in a pattern that will overlap the blasted crater radius of each hole. The optimum charge for this type of blasting is achieved through the use of explosives that have high energy and high detonation velocity. Despite its low cost, the use of ANFO has not proved practical due to its relative low energy level and lack of water resistance. --- ## BENCH BLAST DELAY PATTERN The drill patterns and blasting products used in benching underground are similar to those used in surface mines and quarries, except that a smooth wall blasting technique such as presplitting is normally used to reduce over break and loose rock on the remaining pillar walls. Most underground bench blasting delay patterns utilize a double echelon delay sequence as illustrated in figure 35.38. ![Figure 35.38 – Typical bench blasting pattern and echelon. (Adapted from ISEE Blasters' Handbook™, 17th Ed. figure 25.32)](images/743.png) --- ## BLASTING PRACTICE IN UNDERGROUND COAL MINES Underground coal mines that utilize drilling and blasting to mine the coal are most commonly developed as a room and pillar arrangement. A series of four to six parallel entries typically on 30.5 meter (100 foot) centers with crosscuts connecting the entries create the access. The entries are driven a predetermined length and, depending on conditions, the pillars of coal will later be sub-divided with additional crosscuts in a retreat system to safely recover as much coal as possible as the mine is depleted. --- ### General Discussion There are two general methods of blasting coal in underground mines. These methods are shooting "off the solid," where the only relief is the face that is being advanced, and machine cut "kerf" headings. The diverse geology of bituminous coal seams results in a wide variability of seam thickness, type and amount of impurities, cleavages, brittleness, and characteristics of the rock formations that make the roof and floor of the mine. These factors greatly influence the blasting methods and types of explosives required. Mining regulations require that only a special type of explosive and detonator be used so to minimize the risk of initiating a methane gas or coal dust explosion. These explosives emit pass a defined series of tests. These tests include chemical composition, fume characteristics, gap sensitivity, penetrability for the explosive to ignite methane and / or coal dust in gallery firing; and other tests that will classify the mine as "Permissible" or "Approved" for that type of mining. Formulating explosives so as to minimize the production of mine-hazardous particles and flame on detonation is generally achieved by adding salt to nitroglycerin-based explosives and water to emulsion and water gel explosives. The maximum diameter of the cartridge is limited by the air overpressure that the environment requires that only electric detonators listed with nonincendive delay periods be used with approved products in underground coal mining. Specific safety requirements are referenced in all related Mining Regulations, MSHA, and NFPA as guiding material shells. In the United States the blaster must obtain an exemption or variance from governmental mining regulators to use detonators of any other type of construction. The drilling equipment and the types of bits used to drill the desired borehole diameter all affect the cost of blast. Drilling coal with conventional rotary bits can cause a buildup of heat around each hole; too much burden can result in either inadequate breakage or a blown out shot hole. Insufficient burden will result in excessive fragmentation of the coal and flying debris. Both under and over breakage are likely to cause an ignition of methane gas or coal dust. --- ### Explosives Loading and Stemming Normally, loading to half of the borehole length with explosives will be sufficient to produce good fragmentation with maximum energy in the back of the borehole to shear the solid coal to the cut. Additional explosives "strung out" to the collar of the borehole usually represent wasted energy. All drill cuttings should be removed from the boreholes before loading. Leaving drill cuttings inside the borehole makes it difficult to load the cartridges to the borehole bottom and can also lead to misfires. The recommended loading method is to place the complete charge, including primer, into the collar of each hole in the form of a continuous "string" of cartridges that is pushed to the bottom of each hole in one complete charge. "String" loading will minimize column separation caused by coal dust collecting between cartridges. Stemming plugs can then be tamped into place on top of each charge. U.S. Federal regulations limit the amount of explosives that can be loaded per borehole and require the use of a minimum amount of noncombustible stemming material. An exception is made when blasting solid rock in its natural deposit and for use in anthracite mines for "chute blasting" or for blasting a coal overhang. --- ### Misfired Explosives Deflagrations, dead pressing (See chapter 11) and deformed cartridges are well known problems with explosives in soft minerals like coal. These conditions usually occur as a result of not "string" loading the cartridges as a continuous charge or when an explosive charge is disrupted or displaced by a preceding detonation of an adjacent blast hole. If this condition occurs on a regular basis, the procedures in table 35.8 should be adopted to reduce the risk. ### Strategies To Reduce Risk Of Misfires In Underground Coal Blasting | Procedure | Comment | |-----------|---------| | String load cartridges | This helps prevent air gaps or coal dust cutting between cartridges. | | Adjust delay interval between boreholes. | Reduce delay interval between adjacent boreholes to less than 70 milliseconds | | Adjust borehole/charge length | Increase borehole/charge distance to a minimum of 60 centimeters (24 inches) | *Table 35.8 – Strategies to reduce risk of misfires in underground coal blasting.* --- ### Blast Sequence Each shot in coal mining terminology is an individual borehole detonation in the round initiated in sequence from the opening borehole. Generally, U.S. Federal regulations mandate that the opening borehole must have an in-hole delay of at least 25 milliseconds. This practice is to allow the electrical firing pulse to drop to zero in the firing circuit before the first charge fires. The interval between the designated delay periods of successive shots shall be at least 50 milliseconds, but not more than 100 milliseconds, with the total delay on the entire round not to exceed 1,000 milliseconds. The borehole nearest to a relief cut should be fired first so as to break and displace the coal and provide room for holes that will fire later. Only a "permitted" or "permissible" blasting machine must be used to initiate the electric detonators. In the United States, no electrical device that has not been rated "permitted" or "permissible" for blasting may be used in a coal mine except where "variance" approval is given by the authority having jurisdiction. --- ### Shooting "Off the Solid" Where kerf cutting machines are not available, the "cut" must be blasted. The most common method is the V-cut as is shown in the figure 35.39. A short borehole, approximately 1.8 meters (6 feet) deep, is drilled in a "V" or wedge configuration with the back of the holes as close in line as possible. The boreholes across the face are drilled to the full depth of the proposed round in a fan pattern out to the rib. ![Figure 35.39 – "V" coal drill pattern and delay sequence.](images/745.png) Arranging the delay sequence to the ribs will pile the blasted coal toward the middle of the entry resulting in less damage to the support timbers and facilitating easier loading of the coal. The other type of blasting pattern used for shooting "off the solid" is a "fan" pattern as shown in figure 35.40. A short hole is angled at 45° to the face at either rib, with the successive boreholes increasing in angle until the holes are perpendicular to the face. The pattern should alternate side-to-side with each round. When "shooting off the solid" the explosives ratio (powder factor) will normally be much higher than when blasting to a "kerf". ![Figure 35.40 – Fan underground coal pattern and delay sequence.](images/746.png) --- ### Shooting to a Kerf A kerf provides additional full hole depth relief on two or more faces for a solid face of coal to be broken and permits a more uniform distribution of the total explosives energy. A top or bottom or a combination of kerf cuts is shown in figures 35.41 through 35.44. To achieve maximum benefit from the kerf cut, it is essential that all cuttings be removed to provide additional void space for expansion of the blasted coal. A floor kerf filled with water will result in the same effect as one that is not cleaned out. If the water cannot be removed, additional drill holes and explosives will be necessary to produce adequate results. ![Figure 35.41 – Underground coal blast with undercut and right shear kerf with delay sequence.](images/747.png) ![Figure 35.42 – Underground coal blast with top and right shear kerf with delay sequence.](images/747.png) ![Figure 35.43 – Underground coal shot with horizontal center kerf with delay sequence.](images/747.png)