72
Blast Design for Underground Mining Applications
Roger Holmberg,* William Hustrulid,* and Claude Cunningham*
INTRODUCTION Good blast design and execution are essential ingredients for successful underground mining. Poor blasting practices can have a severely negative impact on the economics of mining. Military blasting, rather than precision blasting, can result in overbreak and dilution of high-grade ores. Military blasting can damage sensitive or tender rock structures that make up the hangingwall or footwall so unwanted caving occurs with the possibility/probability of ore loss and/or dilution. Poor design and execution when ring drilling can mean that succeeding rings are damaged and unchargeable. Failure to complete undercuts can mean the transmission of very high loads to the underlying structures and their subsequent failure. The results are lost revenues and added costs. 1 2 3 Good drilling and good blasting go hand in hand. If the drilling is poor, there is little one can do to correct the rest of the job. It is similar to building a house. If the foundation is poorly done, there are major problems along the rest of the way. A discussion of drilling practices is beyond the scope of this chapter, but the blast design must begin there. Prior to designing the blasting, one must be sure that the miners have the machines and the capabilities to build the design. If the design cannot be built with the tools at hand, then it is no design. Hence, one starts the design process by carefully examining the drilling capabilities. Fortunately, great improvements have been made in machine-
Fortunately, great improvements have been made in machinebased drilling precision capabilities over the past few years. These include laser alignment, using tubular steel, in-hole guidance, and boom alignment instruments and techniques. However, the miners must then use the capabilities of these improved machines to the fullest. Although a good blast design base does exist for underground mining applications, it has often been poorly explained and documented in the literature. This chapter tries to correct this situation, at least to some degree, by offering some initial design guidance for: Bench blasting Ring drilling Crater (VCR) blasting Drifting In this short chapter it is impossible to provide full coverage of such a broad topic as underground blast design. Several different approaches are described, and examples are given as appropriate. A number of simplifications have been made to facilitate the presentation. This is to be regarded as providing an introduction to the topic and not as a cookbook. The reader is advised to contact an explosive supplier and/or other specialists for help in solving complex blasting problems. • • • •
EXPLOSIVE PROPERTIES There are a large number of properties that one considers when choosing an explosive. In terms of doing a blast design, however, the properties are relatively few. They can be simplified to: Density Weight strength Specific gas volume For the rock, the properties are also relatively few: • • •
Density Fracture conditions Rock strength, modulus, and toughness Water conditions Figure 72.1 shows the process involved in developing a blast design. One starts with the results that one wants to achieve and then works to the final design. These results may be • • • •
and then works to the final design. These results may be expressed in the form of a certain desired fragmentation, minimum blast damage to the surrounding rock, vibration limits of sensitive structures, and/or minimum overbreak. Here, for simplicity, it will be assumed that the explosive-rock interaction can be defined in terms of a "powder factor." Although the term "powder factor" is widely used in mining, it must be used with care. It may be expressed in terms of the amount of explosive required to fragment either a certain volume or mass of rock. Thus, it is sometimes expressed in the units of kg/m3 or kg/ton. The difference between these is the density of the rock. For the same generic rock type, the required powder factor to yield the same fragmentation may be quite different, depending on the initial fracturing condition. When discussing powder factor, one must also indicate the explosive that has been used, because the breaking characteristics depend upon the characteristics of the particular explosive in that rock. For example, one kilogram of ANFO has a different energy content than does one kilogram of an emulsion explosive. Even for explosives having the same energy content as expressed in kcal/kg or MJ/kg, the distribution of the energy in terms of the shock energy and the gas expansion energy can be rather different. This partitioning depends upon the characteristics of the rock mass as well as the explosive, so the process is complicated. In performing an actual design, one might consider several different explosives. Then it is important to have a way to examine their relative performances and associated costs on paper. Langefors and Kihlström (1963) suggested that, when comparing the relative strengths of explosives, one should consider both the weight strength and the gas volumes produced. The relative strength of explosive A with respect to ANFO is expressed by:
SANFO = 5 QA/(6Q) + VGA/(6VG) where SANFO = relative strength of explosive A with respect to ANFO QA = weight strength of explosive A (kcal/kg) Q = weight strength of ANFO (kcal/kg) VGA = gas volume of 1 kg of explosive A at STP (l) VG = gas volume of 1 kg of ANFO at STP (l)
The values of weight strength are often given in the specification sheets for different explosives. The values for the gas volumes are seldom given; hence, one is often forced to use the simplification that the relative weight strength of explosive A compared to ANFO is
Unfortunately, in the literature there is no standard value for Q. Here, to provide an index, Q will be assumed to be 912 kcal/kg or 3.82 MJ/kg. The bulk strength (BS) is the weight strength times the density and has the units of kcal/cm3, MJ/cm3 or
times the density and has the units of kcal/cm3, MJ/cm3 or equivalent. The relative bulk strength of explosive A having density &##961;A compared to ANFO having density &##961; is:
where BSANFO = relative bulk strength of explosive A compared to ANFO &##961;A = density of explosive A &##961; = density of ANFO
As an example of the application, consider &##961; = 800 kg/m3 &##961;A = 1200 kg/m3 Q = 912 kcal/kg QA = 850 kcal/kg
The weight strength of explosive A relative to ANFO is:
The bulk strength of explosive A relative to ANFO is:
Because it is a particular hole volume that is to be filled with explosive, the bulk strength is of major importance when considering the blast design. From a drilling point of view one must look at the ability to drillout the design. The reasons for the problems are: Collaring inaccuracies Set-up errors In-hole deviation As will be seen, these problems affect all of the different operations from blasting the cut in a drifting round to stoping patterns. • • •
BENCH-BLAST DESIGNS Introduction Bench-type blast designs are used in a number of different underground mining applications. They include:
underground mining applications. They include: Stoping with parallel holes (Figures 72.2 and 72.3) Drifting in overhand cut-and-fill (Figure 72.4) Up-holes in shrinkage stoping Benching in room-and-pillar mining (horizontal or vertical) (Figure 72.5) • The stoping holes in drifting Normally a square pattern of holes (rows and columns aligned) rather than a staggered pattern is used (Figure 72.6) because the final opening shape is normally square. The hole pattern itself can be rectangular or square. There are two approaches to arriving at a design. Both start with knowing the hole diameter (D) to be drilled. • • • •
The Ash Approach
A standard approach to designing open-pit blasts is based upon the design guidelines of Ash (1963). The same approach, with some minor modifications, can be applied to bench-type blasting in underground mining. It can be shown that the burden B, the distance from the hole to the closest free surface, is directly related to the diameter of the blasthole (D). This is expressed as: B = KB D The hole spacing (S) is related to the burden by: S = KS B and the stemming (T) can also be related to the burden through: T = KT B When examining actual operations, Ash found that for ANFO used in rocks having a density of about 2.5 g/cm3:
KB = 25 The design-spacing factor KS was in the range of (1 to 2) B. The practical range for open-pit mines is from 1 to 1.3, based on energy distribution considerations. The average stemming factor is 0.7 B. Hustrulid (1999) has shown that the factor KB can be related to other explosives using the relative bulk strength relationship: KB (explosive A) = KB (ANFO) (BSA)1/2 For the example given above: KB(explosive A) = 25(1.4)1/2 = 29.6 The value used for KB also depends upon the density of the rock mass. For rock densities significantly above or below 2.5 g/cm3, this factor must be considered. For underground applications, the spacing ratio is about the same as used for open pit mining: KS = 1 to 1.3 Assume that in an open-pit mine, that holes with a diameter of 4 in (100 mm) are used with ANFO (density of 0.8 g/cm3) in a rock with a density 2.5 g/cm3. For simplicity, it will be assumed
rock with a density 2.5 g/cm3. For simplicity, it will be assumed that: KB = 25 KS = 1.3 Using these together with the hole diameter, the burden and spacing become: B = 2.5 m S = 3.25 m The explosive charge/meter of hole (Mc) is given by: Mc = &##960;/4 D2 &##961;ANFO = &##960;/4 (0.100)2 800 = 6.28 kg/m The volume of rock broken per meter of hole is: V = B &##8734; S &##8734; 1 = (2.5)(3.25)(1) = 8.125 m3/m The amount of rock broken per meter of hole is given by: W = SGrock &##8734; V = 2.5 &##8734; 8.125 = 20.3 tonnes/m where W = mass broken/m of hole V = volume of rock broken/m of hole SGrock = specific gravity of rock.
The powder factor (ANFO) expressed as required charge per mass then becomes: PF(ANFO) = W/L = 6.28/20.3 = 0.31 kg/tonne The use of (ANFO) in the expression simply means that it applies for ANFO. In terms of volume, the powder factor (ANFO) becomes: K(ANFO) = L/V = 6.28/8.125 = 0.77 kg/m3 This is very similar to the typical ANFO powder factors used in open-pit mining. However, in open-pit mining the loading equipment is often large and there is lots of room to handle oversize. The same is not true underground, where the presence of oversize can cause significant handling problems. To
of oversize can cause significant handling problems. To overcome this, the powder factor is increased to produce a finer average fragmentation. The authors have found that this can be taken into account by simply reducing the KB values in the Ash formulas when applying them underground. The following value is used as a first approximation: KB = 20 for ANFO In rock with a density of 2.5g/cm3. If the above example is repeated with this value one finds that: B = 20 (0.1) = 2 m S = 1.3 B = 2.6 m V = 5.2 m3/m W = 13 tonnes/m
and the powder factor becomes: PF (ANFO) = 6.28/13 = 0.48 kg/tonne As will be seen later, this is very similar to that experienced in practice. The Powder Factor Approach In the powder factor approach, one begins with a given hole diameter and assumes (or knows) the required powder factor. One can then write the volume of rock that is broken for a fully charged hole of length L with a burden B and spacing S. This is: V = B &##8734; S &##8734; L The mass of rock broken is: W = B &##8734; S &##8734; L &##8734; SGrock where L = charged length SGrock = the specific gravity of the rock The amount of explosive in the hole is given by:
Vexpl = &##960;/4 D2 &##961;expl Using the known powder factor PF, one can combine equations (72.14) and (72.15) to yield:
Assuming that the hole diameter is 4 in (100 mm), the explosive is ANFO with a density of 800 kg/m3, and the required powder factor is 0.48 kg/tonne. Substituting these values into Equation (72.16), one finds that:
Now one must choose either B or S to solve for the other. Assume that: S = 1.3 B Equation (16) then becomes: 1.3 B2 = 5.236 m2 Solving for B, one finds that: B = 2.0 m Thus, the pattern would be: B = 2.0 m S = 2.6 m and the powder factor is that required. Fragmentation The fragmentation depends somewhat on the burden-spacing ratio, and the same powder factor will give different fragmentation depending upon the rock-mass condition and the explosive used. The timing is extremely important to the final results achieved. However, these two approaches are a way to begin the design. For further information, the interested reader is referred to the papers by Cunningham (1992) and the books by Persson, Holmberg and Lee (1994) and by Hustrulid (1999). Wall Damage
Along the walls, special designs are used to minimize unwanted damage. This procedure will be discussed under the drifting section. Data from Canadian Mines In the Canadian Mining Journal's "2000 Mining Sourcebook," a large amount of data collected from Canadian underground mining operations is presented. The blast-pattern data for the open-stoping and blasthole-stoping operations are summarized in Table 72.1. One can see that a wide range of hole diameters and explosive types are included. The spacing:burden ratio varies from 1 to 3.5, with the most common values in the range of 1 to 1.3. Figure 72.7 is a plot of burden versus the hole diameter. Due to a lack of complete information, the different explosive types were not taken into account. Superimposed on the data set are lines corresponding to different KB ratios. As can be seen, the data fall between the lines of KB = 15 and KB = 30. The spread is natural considering the different explosives and rock densities involved. Assuming that KB = 20 as a first approximation is probably good enough for starting the design. Based on the results, the pattern can then be spread or drawn in. Table 72.2 is a summary of the powder factors and the host rock types involved with the different operations. The values range from 0.27 to 1.05 kg/tonne.
Summary As a first approximation, the design approach used in surface mining can be applied to bench-blasting geometries underground. The principal change involves increasing the powder factor to provide the required finer fragmentation.
RING BLASTING DESIGN
Introduction Ring blasting is fundamentally different from all other forms of blasting because the holes are drilled from a central point in and radiate outwards to the limits of the ore block being mined. As shown in Figure 72.8, the technique involves three distinct operations:
Driving a tunnel (or tunnels) along the axis of the proposed excavation. This is the "ring drive." • Creating a vertical slot at the end of the ring drive(s) to the full width of the excavation. • Drilling sets of radial holes called "rings" parallel to the slot. These rings are then blasted progressively into the slot. Figures 72.9-72.13 show some typical designs and demonstrate the basic benefits of the technology. •
The method is very flexible in coping with variations in block size and shape. • It enables large blocks of ground to be blasted from a few access points, thereby improving both development costs and ground stability. • It offers high productivity and insurance against breakdowns because the holes can be drilled and charged well in advance of blasting operations. •
• It is safe, as men do not enter the stope. In fact, as underground mines seek to meet the challenges posed by the new century, many are seeking to maximize the benefits of ring drilling by harnessing modern drilling technology to extend its application and achieve previously undreamed-of efficiencies. Unfortunately, there is also a downside to ring blasting
Because the blastholes radiate from a central drilling position, the coverage of the block of ground varies from minimum at the toe position to maximum at the collar position. This leads to continuously varying powder factors, greater explosive consumption than is theoretically necessary, and complications in charging the holes so as to minimize overcharging. • The holes are drilled at different angles, leading to varying deflection forces for the holes and a potential for gross discrepancies between the designed and achieved drilling patterns. This can lead to poor blasting results and failure to reap the potential production from each ring. • The holes meet the block outline at different angles, making it difficult to ensure that an appropriate and equitable concentration of explosives is achieved in the •
equitable concentration of explosives is achieved in the critical breaking area at the toe. • The holes are drilled at different angles and to different lengths, being collared at close intervals. This imposes a strain on the production team in terms of achieving the planned layout. In addition, if in-hole initiators are used, stock controls over lengths and delays can be difficult. • Up-holes are normally used; these present added costs and problems in terms of retaining explosive, especially where large-diameter holes are concerned. • Ring holes are very prone to misfires caused by cutoffs in the collar area, where the holes are closely bunched and excess energy tends to be available. On the other hand, inter-hole delays are necessary, especially with large blastholes, to reduce the vibration levels. Therefore, selecting and implementing the initiating system is critical. The extent to which the above challenges can be met depends both on an understanding of the extent of the problem posed, and on ways in which good design can reduce or eliminate each problem. The major recurring problems related to blast design is as follows: Failure of up-holes to break out at the toe, leaving `crown' pillars • Poor drawpoint availability caused by the enormous boulders created by back-damage, allowing blocks behind the ring to fall into the drawpoint • Excessive secondary breaking due to a poor distribution of explosives related to both design and implementation problems. In this section, an approach to ring design is presented that will implement intrinsically productive designs, or at least give advance warning of problem areas. •
Basic Mine Layout There is a general desire to increase the rate of bringing a mine into production at minimum cost by having very great vertical separation between the mining levels. This means that rings need to be drilled so as to reach between levels, and the longest holes will be those that just reach to the furthermost corner of a block. First, serious consideration needs to be given to the interlevel spacing because the longer the holes, the greater will be their deviation, and the greater the problem if the drill string
their deviation, and the greater the problem if the drill string should get stuck during drilling, or the ho1e should become blocked prior to blasting. In addition, very large ore blocks will call for more holes to be drilled from the same position, multiplying the bunching effect around the ring drive. It is difficult to believe that angular increments of fractions of a degree are meaningful in this context, yet this is what is called for if reasonably uniform drilling patterns are required. Ideally, the holes should not be more than 25% from their planned positions, let alone intersecting each other. Table 72.3 gives different lengths for holes designed to have a spacing of 3.5 m and an angular deviation of 25%, owing to both setup and wander.
Table 72.3 shows that, above 40 m, exceptional drilling equipment, setup procedures, and operator commitment is required. A further consideration is that very large rings reduce the ability of a mine to source production from different areas in the interests of grade, or of keeping production up in the event of problems with the production ring. The decision as to the relative areas to be assigned to up-rings as opposed to down-rings needs to be seriously considered. Upholes should be drilled as short as possible, because their cost and the work involved in charging them is greater than for down-holes, and the differential increases with length. In addition, the ability to load to a controlled collar distance is much greater with down holes than with up-holes. Apart from operational factors, spillage of explosive from up-holes is inevitable, the extent depending on the borehole conditions, the skill in the handling of loading equipment, and the variability of the explosive being use. Waste reduces efficiencies, but the safe and hygienic disposal of spilled explosive also impacts on the overall productivity. Thus, while it is clear that the effective loading of long up-holes is technically possible, it is highly desirable to facilitate charging operations by locating the ring drive high in the ore block. If the ring drive is no more than 15 m below the top of the block, loading operations are much enhanced. The shape of the ore blocks--especially their
enhanced. The shape of the ore blocks--especially their placement opposite one another and other profiles--is extremely important. It is easy to create "shadow" areas inaccessible by drilling from the available ring drives, or areas that result in an uneconomic drilling pattern owing to indentations or "nodules" at critical points. These may be unavoidable, but can sometimes be prevented during the design stage. General Design Principles Good fragmentation can result only if sufficient explosive is used in correctly drilled holes, with initiation arranged to ensure that all the explosive detonates in the proper sequence. Owing to the complexity of ideal layouts for ring blasting, however, compromises have to be reached between what is theoretically desirable and what can reasonably be implemented. Blasthole Length. When laying out rings, one of the main considerations is to limit the maximum hole length. To minimize development, holes are often laid out over excessive lengths simply to span the distance between sublevels. This creates the following problems: 1. Inaccurate drilling--the small-diameter holes used in this technique can seldom be drilled with reasonable accuracy over more than about 20 meters with conventional equipment. Beyond this length, and especially with angled holes, severe departures from the planned burden and spacing are likely, resulting in poor fragmentation, toe formation, and overbreak. 2. Gauge reduction--used as the hole results in reduced critical toe area. Smaller drill bits are often used as the length increases, with explosive loading in the critical toe region. 3. Drilling inefficiency--the efficiency of blow transmission in the drill string decreases by about 10% across each extension coupling and this results in reduced penetration rates and increased equipment wear over long holes. In ring blasting, the situation is even worse because of the short extension steels necessarily used in constricted ring drives. Ideally, both up-and down-holes should be drilled to double the distance between sublevels, but in sublevel caving, for example, only up-holes can be drilled. This factor should be taken into
only up-holes can be drilled. This factor should be taken into account when laying out sublevels. As a rule, the majority of holes should not exceed 30 meters unless specialized equipment (operated by properly trained drilling crews) is used. With longer holes, poor fragmentation is likely to be inherent to the system and can only be compensated for by adopting a higher explosive-loading ratio than would normally be considered. Determining Drilling Pattern. The terms "spacing" and "burden" require definition because of a lack of general agreement among ring-blasting operators. "Burden" means the distance between two consecutive rings. "Spacing" means the distance between the ends (or "toes") of neighboring holes in one ring, measured at right angles and straddling the outline of the ore block, using construction lines (see Figure 72.14). The correct drilling pattern is one that delivers the appropriate energy at the toes of the blastholes. If it can be shown that a particular drilling pattern in a parallel-hole benching operation gives the required blasting results, this pattern should be adopted for ring blasting. An example of this is tests that were performed in kimberlite. The ring-drilling pattern was derived from preliminary blasting trials in the last months of the open-pit operation, with fragmentation as the criterion. In the absence of such guides to hole spacing, a powder factor approach can be taken. Care must be exercised, however, because it does not allow for special geological conditions, specific breaking results, the blasting geometry, or the effect of explosive type.
The following method of calculating burden and spacing ensures good breaking in the critical toe area around the ends of the
good breaking in the critical toe area around the ends of the blastholes. For simplicity, the calculation assumes that all holes are parallel, with the explosive column reaching to a collar length of twenty charge diameters. The hole length used in the calculation is the average length in the ring. For ANFO or slurry, the explosive diameter equals the hole diameter. The fact that the powder factor increases towards the ring drive through convergence of the holes is compensated for in practice by using different uncharged lengths in each hole. Alternatively, spaced charges or charges of lower strength can be used, but these solutions are not recommended because of the complications they introduce. The formula relating the burden and spacing to the powder factor is:
where B = nominal ring burden (m) S = nominal toe spacing (m) L = length of explosive column (m) H = average hole length for ring drilling (m) Mc = explosive mass per unit length (kg/m) K = powder factor (kg/m3)
For simplicity, we will assume that all holes are parallel, and the uncharged portion is equal to 20 charge diameters. The charged hole length becomes: L = H - 20D where D = explosive diameter (m)
In general, the toe spacing of holes should exceed the burden on the ring, but the exact ratio of spacing to burden is not critical. The spacing:burden ratio is normally assumed to be 1.3, but can be as high as 1.5. In normal ring blasting, good fragmentation is required and fairly high powder factors are necessary, whatever the ground hardness. A starting figure of 0.8 kg/m3 would suit most types of weak rock, while 1.2 kg/m3 would suit a dense, strong rock. The design powder factor depends on the toughness and blockiness of the ground and whether or not tight breaking conditions prevail. It is good practice to allow for at least 30% expansion from the solid when blasting into a restricted slot.
expansion from the solid when blasting into a restricted slot. A major error to be avoided is using a planned powder factor across the whole ring. This approach is likely to result in excessive hole spacing, since convergence of the holes always causes overcharging inside the pattern. It is patently poor blasting practice to tolerate inadequate energy at the toe portion just so an average powder-factor can be obtained. To indicate how the calculation process works, consider using an emulsion explosive with a density of 1,180 kg/m3 in a 102-mmdiameter blasthole. The powder factor is 0.8 kg/m3. For simplicity, we will assume that the explosive extends to the collar and that the hole length does not have to be considered. In Equation 72.17: Mc = 9.64 kg/m K = 0.8 kg/m3 and thus
Assuming that: S = 1.3 B then 1.3 B2 = 12.05 and B = 3.05 m Therefore, the "nominal" drilling pattern is 3.0 m &##8734; 4.0 m. The rings should, therefore, be drilled 3 m apart along the ring drive and the holes angled to terminate with their toes 4 m apart. Laying Out the Ring. Once the "nominal" drilling pattern has been obtained, it must be applied to the block outlines. The
been obtained, it must be applied to the block outlines. The spacing is measured perpendicular to the blasthole. Because most of the holes will not meet the block outline at right angles, the holes have to be fitted so that no energy starvation takes place between them along the oblique parts of the outline. This is done by plotting the hole positions beyond the outline and imposing the designated spacing between the holes so as to straddle the outline equally (Figure 72.14). Each variation in the ore block requires a new layout. The exercise is easily and quickly performed by hand as follows:
1. Holes are drawn in from the drilling position to the corners of the ore block. 2. Using a scale and 90&##176; offset, the toes of intervening holes are marked off consecutively, using the calculated spacing, as shown in Figure 72.14. 3. As the intermediate holes approach fixed corner holes, small adjustments are made. The hole before the corner hole is (a) either centralized between the corner hole and the next one back or (b) omitted. This is done at the discretion of the person doing the layout. When it is omitted, the next hole back should be centralized, if necessary. This practice rarely results in more than 10% deviation from planned spacing. When this process is complete, the last hole will very seldom fit exactly on the last part of the outline, and the spacing should be readjusted, reducing or increasing the spacing slightly until all the holes are equally spaced
spacing slightly until all the holes are equally spaced (Figure 72.15). 4. On completion of the ring layout, the actual angles and lengths of the holes are measured and entered on the instruction sheet (Figure 72.16).
If the ring is not vertical, the exercise must be done in the plane of the ring. Charging Pattern. The blast holes converge towards the ring drive and, to avoid serious overcharging, require an alternating pattern of uncharged collar lengths. This pattern has to be simple and repetitive if charging crews are to adhere to it meaningfully, and the following compromise arrangement is suggested. Assume three uncharged collar lengths: Ts = 20 explosive diameters Tm = 50 explosive diameters Tl = 125 explosive diameters.
The holes are numbered beginning at a readily recognizable hole, e.g., the lower left-hand side. The uncharged lengths are then specified in that order:
then specified in that order: Ts, Tm, Ts, Tm, Ts, Tl, Tm, Tl, Ts, Tl, Tm, Ts, Tm, Ts, Tm, Ts Prior to charging up, a responsible person attaches tags indicating whether Ts, Tm, or Tl apply to the collar length for each hole. It should be noted that no uncharged collar should be more than two-thirds of the hole. If Tm exceeds two-thirds of the hole length, use Ts. If only Tl exceeds two-thirds of the hole length, alternate Ts and Tm until holes longer than 1.5 Tl are encountered. The end holes (which only occur where full rings are not drilled) are always given Ts. When the planned powder factor and the actual projected explosives used per cubic meter broken are finally compared, there may be a significant discrepancy, caused by using approximations to specify the uncharged lengths. This is unavoidable without either sacrificing fragmentation or imposing a more complex charging system. It can be justified on the grounds that a higher powder factor will result in finer overall fragmentation, which is usually welcome in ring-blasting operations. In trying to reduce the eventual overall powder factor, it is important not to decrease the powder factor in the original calculation. This would result in poor breaking in the critical region of the ring perimeter. If a large number of ring designs must be made, the process is very tedious and time-consuming if done by hand. Software has been developed to perform this task. The software "Ring" developed by AECI's Blast Consult Group is one such example. It provides optimized ring-drilling designs in a minimum of time. This is extremely useful in comparing the effects of different hole diameters and checking the achieved hole spacing against the design. The package normally uses the nominal spacing as a base to find the number of holes in the ring and the actual spacing. It can also be given the required number of holes and can derive the spacing from this. Once the correct outline pattern is achieved, the spacing can be offset against the designed powder factor to correct for excessive or inadequate energy in the toe area. For example, if the exercise in the above example yielded a final average toe spacing of 3.7 m, instead of 4.0 m, Equation 72.17 cou1d be used, which would derive a ring burden of 3.25 m instead of 3.0 m. Optimizing Results
The basic approach to designing the layout of ring-blast holes and distributing explosives within each ring was discussed in the previous section. This section shows how blasting results can be optimized by suitable choice of hole diameter, explosive type, and initiation system. Priming and initiation of holes is also discussed. Choice of Hole Diameter. Hole diameter and explosive density basically determine the hole spacing. An increase in diameter increases the spacing and enables fewer holes to be drilled. This means that large-diameter holes are a good choice for large ore blocks. The largest feasible hole diameter should be used for ring drilling. This has the following benefits: Less drilling--Large holes take more explosive, which can break to wider patterns. This means the holes are less crowded around the ring drive and charging operations are faster, and greater production is obtained from the available equipment. • Easier loading--Larger holes tend to suffer less from obstructions that prevent cartridges (or loading hoses) from being pushed to the back of the hole. Using hole diameters below 45 mm may necessitate charging with 32 &##8734; 200-mm cartridges, which have the added disadvantage of being shorter and taking longer to load. The largest diameter that should be considered for up-holes is about 100 mm. Above this, retention of the explosive is a major problem. •
It often happens, however, that the same drilling equipment used for large blocks is also used for small blocks, e.g., production cones. It is patently inappropriate to use a hole diameter that calls for a hole spacing of 3.5 m if the available space is 4.5 m, or even 9.0 m. Each available block accommodates a particular number of equally spaced holes and the smaller the block, the less the ability of large holes to deliver the ideal hole spacing. This can be particularly so when the outline includes "notches" because of the intrusion of other openings or geological features. A further consideration is the extent to which up-holes will be required. Holes in excess of 64 mm are difficult to load with
required. Holes in excess of 64 mm are difficult to load with cartridged explosives, while those in excess of 89 mm require special skills with ANFO. For holes in excess of 115 mm, the effort needed to retain any explosive is so great that only short, relatively infrequent holes should be considered. If rock deterioration is a serious concern, it should be remembered that large-diameter holes deliver greater vibration energy to the rock. The explosive energy can be reduced to counter this by using smaller-diameter holes. Selecting Explosive. From the foregoing, it will be apparent that selecting the explosive is likely to follow naturally from the choice of hole diameter, which itself is largely determined by considerations of ore-block size and the extent to which upholes are to be drilled. Almost any explosive is suitable for nearvertical down-holes, but an in-hole transport system is necessary as the inclination to the horizontal decreases below 50&##176;. The same restrictions apply for up-holes. In larger holes, ANFO and pumped emulsion products are the only option; the actual choice will depend on parameters beyond the scope of this paper, not least of which is the availability of suitable loading equipment. High-density, high-energy, well-coupled explosives have the best potential for ensuring good fragmentation in ring blasting. ANFO can only be used in dry blastholes. Its low price and convenience in pneumatic loading, together with its effective breaking, make it a good choice in most conditions, but it is not always used to best advantage. In particular, the ANFO column should be boostered at 5-m intervals to ensure the maintenance of a stable detonation velocity over the entire charge length. Any cap-sensitive cartridged explosive can be used for boosters. Failure to observe proper boostering is bound to result in substandard performance. Charging Holes. Because the blastholes converge, grossly inefficient blasting will occur unless the uncharged collar is varied from hole to hole. What length to 1eave uncharged is not a trivial matter, because it is related not only to the planned hole spacing, but also to the size and shape of the ore block. The primary criterion is that holes should be charged to the point at which the tangential distance from the end of the charge to the next hole is half the designed hole spacing. In
charge to the next hole is half the designed hole spacing. In practice, even skilled charging crews are unlikely to be able to control charging lengths to accuracies better than whole meter lengths, and design considerations should take this into account. When deciding on the uncharged lengths, the software programs are extremely useful. They can quickly determine which regions are under- or over-charged. Initiation. Initiation is a subject of its own. Technically, ideal initiation is basically implemented when each hole is initiated from a point near its toe with very rapid firing of holes spreading from a central point downward on each side. This results in optimized fragmentation while limiting the vibration levels and avoiding cutoffs near the ring drive. Unfortunately, the practical problems with ideal initiation sometimes lead to pragmatic solutions, the chief aim of which is merely to ensure that all the holes fire while the vibration is limited as far as possible. In planning a blast, delay units and primers are allocated to each hole, and judgment must be used as to whether backup units are needed. Ground conditions and loading methods will determine the likelihood of initiation failures, and any sign that a hole may misfire calls for a backup system. Collar priming imposes real limits on timing options because cutoffs automatically result if some holes fire earlier than their neighbors. Delay detonators inside the blastholes are the most commonly used means of initiating ring blasts because they are less likely to be cut off by rock movement and concussion. Continuity of detonation is assured by lining each blasthole with detonating cord. Selection of Delays. Theoretically, a different delay should be used in each hole to improve fragmentation and limit vibration effects, but this has some disadvantages: Limitation of blast size--The range of delays is limited and assigning different delays to each ring would severely restrict the number of rings that could be blasted at one time. Large blasts are usually desirable as they result in fewer oversize slabs of rock • Misfiring of later delays--The collars of the holes converge on the ring drive and with the common practice of collar priming, the detonators are located close together in the rock mass. •
As a result, the first hole to detonate is likely to break out the collars of adjacent blastholes, complete with detonators. This can happen even when detonators are quite deep inside the blastholes, but the likelihood is much reduced if all the delays are the same. In view of these problems, and because it is simpler to implement, it is preferable to specify only one delay per ring. The detonator lead wires should be not less than 4.5-m long, as most cutoffs take place within 3-4 m of the collar, and the detonator should therefore be located at greater depth. Notwithstanding the above, it is sometimes essential to use two or even three different delays per ring to reduce concussion. To minimize the danger of cutoffs, the delays are not alternated between holes, but are apportioned to whole sections of the ring as shown in Figure 72.16. Delay Range. Using alternate rather than consecutive delay numbers is recommended to eliminate the possibility of both out-of-sequence shots and "crowding" between consecutive delays, due to inherent scatter about the nominal firing times of delay detonators. In some cases, large numbers of rings have to be fired in sequence, requiring in excess of the entire range of delay numbers. No attempt should be made to increase the delay coverage by firing pairs of rings on the same delay: choking, overbreak, and poor fragmentation are likely to result. The maximum range of delays is: SPD numbers 1 - 20 LPD numbers 3 - 14 Thus, there is a total of 32 delays. If only some SPD numbers have to be used consecutively in a medium-scale blast, delays 1-6 should be the first ones run consecutively, as there is minimal chance of overlapping with these units. Instantaneous detonators are not recommended for series-inparallel circuits, as their detonation may cause premature dislocation of the blasting circuit, resulting in misfires. Naturally,
dislocation of the blasting circuit, resulting in misfires. Naturally, if only single rings are blasted, these can be primed simply with detonating cord and initiated using an electric detonator or capped fuse. The advent of the electronic delays means that there are many more possibilities for initiating the holes and in the number of rings that can be fired. Priming Ring Holes ANFO. Pneumatic ANFO loading, generates high-voltage static electricity, which can prematurely fire normal electric detonators. Three solutions to this hazard exist as follows: "NONEL" detonators can be used with confidence in pneumatically charged ANFO. The detonator is initiated by a shock wave transmitted through a long, nondestructing plastic tube, which may, however, constitute a contaminant in certain types of ore. • Static-safe electric detonators. These detonators require a 60 MJ/ohm firing impulse (as against 4 MJ/ohm for normal detonators) and can be safely used in pneumatically charged ANFO. They require more powerful exploders and heavier blasting cables. • The simplest, but least satisfactory solution because of the danger of cutoffs is to collar-prime all ring holes with normal electric detonators. The detonators should not be attached to the detonating cord down-line for at least an hour following the completion of loading operations, to allow for the dissipation of any static accumulation. Cartridged Explosives. The detonator is normally fixed inside a primer cartridge, which is then pressed into place using a buffer cartridge or plastic "spider." Where the uncharged collar length exceeds that of the leading wires, the detonator is attached to the detonating cord down line. •
Toe Priming. Because of problems associated with long lead wires, it is not advisable to toe-prime ring holes with electric detonators. This is easily achieved, however, if "NONEL" delay detonator assemblies are used. Blasting Circuit Current leakage from the blasting circuit is a common cause of misfires in ring blasting, especially under the following conditions:
conditions: Where it is wet and conductive salts are leached out of the explosive • In conductive orebodies • Where high resistance iron leading wires are used Therefore, precautions should be taken to maintain effective insulation between the blasting circuit and the country rock. When charging up, the PVC insulation on detonator lead wires is prone to damage from sharp edges or tight kinking. Care must be taken to avoid damage while unraveling the lead wires, making up primers, pushing primers into blastholes, and loading explosive over lead wires. Insulating putty should be used to enclose the bare wires at each connection. Neatly suspended blasting cables are easily inspected and are less vulnerable to accidental damage. An earth-leakage tester is invaluable for checking circuit insulation. Furthermore, all modern detonators are designed to minimize current leakage by operating at a relatively low voltage. •
Conclusion A review of the foregoing indicates that a technically optimized ring-blasting layout can be proposed. The idealized design, then, would be as given in Table 72.4.
While ring blasting techniques offer substantia1 benefits, there are intrinsic problems that are exacerbated when large blocks of ore are mined in attempts to improve the production economics. These problems relate to both the convergence of the holes around the ring drive and the charging of the up-holes. The size
around the ring drive and the charging of the up-holes. The size and shape of the ore block concerned should be careful1y considered so that good drilling patterns can be achieved. The hole diameters chosen both affect, and are affected by, the size of the ore block. The explosive selected depends on the hole diameter and, together with the diameter, determines the nominal drilling pattern. Ring-drilling patterns are best derived through the use of computer software together with engineering judgment. Explosive loading is necessarily high, and charging patterns require careful thought if overcharging is to be minimized and adequate energy is to be available at all parts of the ore block. Initiation systems are critical to the blasting results, but sub-optimal systems must be adopted if collar priming is practiced. In view of all these factors, idealized guidelines can be given for ring design, but it should be recognized that developments in technology might in time change these parameters. A blasting development program that focuses on determining the real effectiveness and productivity of ongoing blasting operations is a worthy investment at any mine that uses ring blasting. The design process for ring blasting is exceptionally demanding, and good software is invaluable for exploring the various options in a timely fashion. Such software is particularly useful if it also provides drilling, charging, and initiation instructions and if it can show the efficiencies and costs for any particular layout.
CRATER-BLASTING THEORY AND APPLICATION Introduction In Canada, a new underground mining method, the Vertical Crater Retreat (VCR) or Vertical Retreat Mining (VRM) method (see Figure 72.17) was developed in 1975 for primary stoping, pillar recovery, and drop raising. This was made possible by the introduction of 165-mm-diameter holes to underground mining.
When vertical (or inclined) large-diameter holes are drilled on a designed pattern from a cut over a stope or pillar to bottom in the back of the undercut, and spherical charges of explosives are placed within these holes at a calculated optimum distance from the back of the undercut and detonated (see Figure 72.18), a vertical thickness of ore will be blasted downwards into the previously mined area. Repeating this loading and blasting procedure, mining of the stope or pillar retreats in the form of horizontal slices in a vertical upwards direction until the top sill is blasted and the mining of a stope or pillar is completed.
The VCR method has been and currently is being practiced in various mines in Canada, the United States, Europe, Central America, and Australia. In this section, the theory of the VCR method will be discussed and then applied to the Luossavaara mine in northern Sweden. Cratering Theory Introduction. The concept of cratering and its development may be attributed to C.W. Livingston. It is a versatile tool for
be attributed to C.W. Livingston. It is a versatile tool for studying the blasting phenomenon, and its application has resulted in the development of a new underground mining method, the VCR method of primary stoping, pillar recovery, and drop raising. A crater-blast is a blast where a spherical or near spherical charge (1:6 diameter-to-length ratio) is detonated beneath a surface that extends laterally in all directions beyond the point where the surrounding material would be affected by the blast. Figure 72.19 shows the nomenclature used in VCR, and they are described below.
&##8709; = Hole diameter 6&##8709; = Charge length db = Depth of burial. Distance from surface to center of charge do = optimum depth of burial. The depth of burial at which the greatest volume of rock is broken N = Critical distance. The depth of burial at which the effects of a cratering charge are just noticeable on the surface r = Radius of crater ro = Radius of crater formed at optimum depth of burial V = Crater volume W = Charge weight
There is a definite relationship between the energy of the explosive and the volume of the material that is affected by the blast. This relationship is significantly affected by the placement of the charge. Livingston determined that a strain-energy relationship exists, as expressed by an empirical equation: N = EW1/3 where N = the critical distance at which breakage of the surface above
the spherical charge does not exceed a specified limit. E = the strain energy factor, a constant for a given explosive-rock combination, W = the weight of the explosive charge
The same equation may be written in the form of: db = &##916; EW 1/3
where db = the distance from the surface to the center of gravity of the charge, i.e. depth of burial &##916; = db/N which is a dimensionless number expressing the ratio of any depth of burial compared to the critical distance
When db is such that the maximum volume of rock is broken to an excellent fragment size, this burial is called the optimum distance: do. For further study of the cratering theory, see Lang (1983). Choosing the Best Explosive for VCR Mining. When the material that is to be blasted remains constant, but several different explosives are considered, the cratering theory may be used to determine the most suitable explosive through the application of Livingston's Breakage Process Equation: V = ABCWE3 where: W = Charge weight V = Crater volume E = Strain energy factor A = Energy utilization number B = Material's behavior index C = Stress distribution number
Inasmuch as V, W, and E can be measured with certainty, it remains for the observer to isolate the variables A, B, and C. The energy utilization number, A, is the ratio of the volume of the crater within limits of complete rupture at any depth, to the volume at optimum depth, where the maximum proportion of the energy of the explosion is utilized in the failure process: A = V/Vo The maximum value of A is equal to 1.0 at optimum depth (where fracturing reaches the most efficient development). Accordingly, the numerical values of A are less than 1.0 at other charge depths.
The material behavior index, B, is a constant for a given type of explosive and weight of charge in a given material. B is measured at optimum depth and: B = Vo /N3 It has been derived from: Vo = B(WE3AC) where A = 1 at optimum depth do C = 1 if the charge is spherical
One can conclude that both A and B describe the effect of the explosive upon the failure process in blasting. The value `A' best describes the effects of the variation in energy density with distance, and B best describes the effects of the variation in energy density accompanying changes in the stress-strain relations as measured at a given reference energy level. The following example will demonstrate the application of the breakage process equation for the comparison of the performance of the explosives in the same rock. Basic cratering research was conducted in a hard, cherty magnetic-iron formation with two types of slurry: Slurry 1 (Selleck 1962) and Slurry 2 (Lang 1962). The curves of V/W versus &##916; for the two experiments are plotted in Figure 72.20. The optimum depth ratio was found to be the same for both explosives: &##916;o = 0.58, but E and N were different.
The values of A were calculated for each crater, and the results were plotted against depth ratio (see Figure 72.21). The two
were plotted against depth ratio (see Figure 72.21). The two curves are similar to those of V/W versus &##916;. This diagram clearly indicates that in the case of Slurry 2, more energy is being utilized in the secondary fragmentation range and in the flyrock range than in the case of Slurry 1. This is responsible for better fragmentation and more gas energy. Production-scale blasts confirmed the results of these cratering experiments.
Material behavior index values for both explosives were calculated at optimum depth and found to be: Slurry 1. Bo = 0.42 Slurry 2. Bo = 0.33 Higher values of B are characteristic of brittle-type failure. Experiments show that B decreases as the material becomes more plastic-acting, which was true in this experiment as well. Slurry 1 had a high detonation velocity; thus the material was acting in the brittle manner. Slurry 2, due to the 10% Al content, had a lower detonation velocity and the load was a slower and more sustained type. Hence, the same rock behaved in a rather plastic manner. The stress distribution number, C, was 1 because both experiments employed spherical charges. One may conclude that when comparing different explosives in the same material, the comparison must be made keeping the geometry of blast constant; otherwise, the results will be misleading.
constant; otherwise, the results will be misleading. 1. Separate cratering experiments should be conducted with the different explosives in the same material. 2. Determine N, &##916;o for each experiment. 3. If this information is for designing VCR-type blasting, then do and optimum spacing should be calculated for each explosive and ore combination. These criteria should be used in each respective stope. Small-Scale Cratering Tests. The purpose of performing smallscale crater tests is to obtain the data required to make qualified predictions of the blasting results in the full production stope. It is necessary to conduct the crater tests as close as possible to the stope where the VCR method will be used. Different rock properties and structural geology may cause an over- or underestimation of the depth of burial for the production blasts. If the depth of burial is less than optimum, it will result in a satisfactory breakage, but the cost for drilling and explosives will be too high. If the depth of burial is larger than the optimum one, bells or unsatisfactory fragmentation may occur. Due to development work in the stope, it is sometimes possible to carry out the test in the stope (in the undercut). Application of Crater Blasting to the Luossavaara Mine Introduction. When the activities in the Luossavaara Research Mine were planned and outlined, it was decided that new mining methods with large-diameter holes should be tested. Discussions during the planning of the Luossavaara Mine resulted in a small test stope D1 (see Figure 72.22) where the VCR method could be evaluated under Swedish conditions. Mr. Leslie Lang, of L.C. Lang & Associates, Inc., in Canada, was engaged as a consultant to the Research Mine and SveDeFo when the test shots for the future design were planned. This section describes the results of the test shots and the proposal that was made for the full-scale production of stope D1.
Small-Scale Crater Tests. In Luossavaara, a field mapping procedure was carried out to investigate whether the abandoned mine area above production stope D would be suitable for crater tests (Mäki 1982). Through comparison to available data, it was found (Mäki 1982 and Röshoff 1981) that "On the basis of the analysis of structure densities, structure lengths and RQD values it may be concluded that no major variation of rock properties exists within the orebody." The data from the test level at 250 m was compared with data from an access drift between the 270- and the 290-m level. However, as the major part of this drift was located in the footwall, it was possible that the conclusion could change somewhat when data concerning the stope itself became available. Cores from profiles in the test area indicated good-to-excellent rock. The RQD values were in the range of 80%-100%. The primary rock types of the side walls were breccia, quartz porphyry, and syenite porphyry. A total of 23 holes were drilled with hole diameters of 38 and 102 mm. The tests were located at the 250-m level above stope D where the rock properties and geological structure were similar to what was expected in the production stope. Eleven 102-mm-diameter holes were drilled and blasted with a non-cap-sensitive TNTslurry, Reolit. Two 102-mm-diameter holes were blasted with ANFO, and six holes with a diameter of 38 mm were drilled and blasted with ANFO. The reason for blasting the 38-mm-diameter holes was to investigate the scale effect. The test holes were horizontally drilled in the rib, perpendicular to the drift. The face was relatively smooth without large hills or valleys. The collars of the 102-mm-diameter holes were drilled 1.6-2.0 m above the floor, and the distances between the holes were not less than 4.5 m. The 38-mm-diameter holes had a spacing of 2.5 m. The length of the holes used for the 102-m-diameter TNT-slurry tests were 3.0, 2.8, 2.6, 2.0, 1.8, 1.6, 1.6, 1.2, 1.1, 1.0, and
tests were 3.0, 2.8, 2.6, 2.0, 1.8, 1.6, 1.6, 1.2, 1.1, 1.0, and 0.85 meters. Only two 102-mm-diameter ANFO test holes were drilled and blasted. One hole was 1.1 and the other hole was 0.9 m deep. In the ANFO test, holes 38 mm in diameter and 1.3, 1.0, 0.8, 0.7, 0.6, and 0. 5 meters 1ong were used. All holes were drilled in the footwall. The holes were flushed clear and measured after drilling. The explosives used in the small crater tests were Reolit, manufactured by Nitro Nobel AB, and ANFO K2Z, manufactured by Kimit AB. To initiate the explosives, a plastic PETN explosive NSP-71, developed by Bofors, was used. Reolit is a non-capsensitive TNT-slurry explosive containing 22 % TNT and 3% Al. The density is 1,450 kg/m3. The weight strength relative to ANFO is 1.20. ANFO K2Z consists of 47.3 % prill, 47.3 % crystalline ammonium nitrate, and 5.4 % fuel oil. The density is 1,000 kg/m3. The charge weight in the 102-mm-diameter holes was 6.7 kg for the TNT slurry and 4.75 kg for ANFO. Using 38mm-diameter holes, the charge weight was 0.25 kg/hole. The 102-mm-diameter charges were primed with 250 g of NSP 71 Bofors and the 38-mm-diameter holes with 25 g of the same explosive. To be able to load the horizontal holes the explosives were packed in 100-mm-diameter plastic bags. To check the performance of the explosive, the velocity of detonation was measured in the 102-mm-diameter holes. The VOD measurement setup consists of probes, a pulse former unit, and a transient recorder, a Nicolet Explorer 1090 A. When the detonation front reaches a probe, the circuit is shorted, and a pulse in the pulse former unit is generated. This pulse has a defined RC constant. A transient recorder registers the positive or negative pulse. By positioning a number of probes along the travel path of the detonation front, a number of pulses, each with its own RC characteristic, can be registered by the transient recorder. By measuring the time lapse between the pulses and by knowing the distances between the probes, the velocity of the detonation wave can be calculated. Three probes for VOD measurements were taped on a PVC rod with a diameter of 5 mm. By pushing the PVC rod into the very bottom of the hole, the distance between the hole bottom and the first probe was fixed. Half the charge was then placed and tamped at the bottom of the hole. The primer was pushed in with a loading stick and with great care tamped close to the inner part of the charge. The remaining part of the charge was finally placed and tamped. The three probes for VOD measurement were now
tamped. The three probes for VOD measurement were now placed in the outer half of the charge. The length from the collar to the charge was measured and a three-part wooden plug kept the charge in position. Stemming consisted of a plug of bentonite and gravel (0-30 mm). A Nonel-cap initiated the primer. After firing the shot, scaling of the crater walls was kept to a minimum. All structural weakness planes, which may have influenced the size or shape of the crater, were noted. Photographs of the craters were taken after each shot. Crater depth as a function of position was determined using a sliding ruler attached perpendicular to a 2.5 &##8734; 2.5-m vertically mounted aluminum frame. From measurements made on a 25cm grid, the maximum depth and radii could be determined and the volume calculated. A total of 11 craters were formed at different depths of burial while keeping the charge weight of Reolit slurry constant: W = 6.8 kg. The results are given in Table 72.5.
The critical distance (N) was determined to be N = 2.5 m, and the calculated strain energy factor (E) was: E = N/W1/3 < 1.32 The plot in Figure 72.23 indicates some scatter. This is due to minor geological discontinuities present in the rock mass. The structural geology will generally have a more overwhelming influence on the cratering results when using relatively small charges than it will with larger charges. The optimum depth ratio (&##916;o) appears to be in the range 0.52-0.6. Additional tests would be needed to reduce the interval. To be on the safe side &##916;o is estimated to be &##916;o = 0.52
in this test. This suggests a predominantly shock type failure of the ore when using Reolit. For &##916;o = 0.52 and N = 2.5 m, the calculated depth of burial for a 6.8 kg charge is:
do = &##916;oN = 0.52 &##8734; 2.5 = 1.3 m From Figure 72.23, the value of V/W corresponding to the optimum depth ratio is V/W = 0.33 and hence V = 0.33 &##8734; 6.8 = 2.24 m3. Representing the crater by a cone with apex at the bottom of the charge, one may calculate the radius ro. ro &##8734; &##960;(do + 6&##960;/2) /3 = 2.24 or ro = 1.2 m The following basic data have been obtained: W = 6.8 kg N = 2.5 m E = 1.32 &##916;o = 0.52 do = 1.3 m ro = 1.2 m
The Production Scale Design. The next step is to scale up these cratering results for a production-scale blast in VCR stopes when the charge weight of the same explosive is increased to W2 = 31 kg and the hole diameter is 6.5 in. Following the Livingston theory, it is assumed that E = 1.32 remains constant. The critical distance for the 31-kg charge weight is:
critical distance for the 31-kg charge weight is: N = EW21/3 < 4.15 The center of this charge should be at the optimum distance Do from the back of the stope: Do = &##916;oN = 0.52 &##8734; 4.15 < 2.2 m The U.S. Army Corps of Engineers, Nuclear Cratering Group, uses another form of the calculation. They calculate a scaling factor (F) of: F= (W2/W)1/3 = (31/6.8)1/3 < 1 .66 The optimum distance for this larger charge is: Do = do F = 1.3 &##8734; 1.66 < 2.2 m The crater radius becomes: Ro = ro F = 1.3 &##8734; 1 .66 < 2.0 m It is important to ensure complete breakage of the rock between two adjacent holes in the stope by designing an optimum spacing between holes. The recommended hole spacing (So) should be in the range of 1.2 Ro to 1.6 Ro. In the case of Reolit it is: Smin = 1.2 &##8734; 2.0 < 2.4 m and Smax = 1.6 &##8734; 2.0 < 3.2 m It is more prudent to design the first stope using Smin and then increase it gradually to Smax. In further stopes, if the results are satisfactory, the pattern can be expanded. However S = 3 m will probably not cause any problem in the Luossavaara type of ore. The advance A will be:
A = Do + 6 &##8709;/2 = 2.2 + 0.5 = 2.7 m The specific charge (q) for So = 3 m is: q = 31/(A &##8734; 32 ) = 1.3 kg/m3 Comments on the Reolit Tests. As can be seen from Figure 72.23 where the smooth curve has been fitted by eye to the experimental points, it is obvious that the curve could have been drawn in various ways. For example, if the curve had been fitted using the least-square method, it would definitely not have looked the same. The interpretation of the results appears to require a considerable amount of subjective assessment on the part of the person conducting the small-scale tests. This means that cratering should preferably be carried out only by persons with previous experience in crater testing and, if possible, production blasting. For instance, the values for shot No.10 were omitted when plotting the curve because the geological mapping indicated that the rock was fractured to a greater extent around this hole than around the rest of the holes. Shot No. 2 had a low VOD, which might explain its small crater volume. If the value for shot No. 10 had been taken as the peak of the curve (yielding the optimum ratio Ao = 0.28), a strong shock-type failure would have been indicated with all its consequences (do, Vo). This would not be consistent with the majority of the information. The ANFO Cratering Results. Tests were carried out with ANFO charges to estimate the breaking capacity with an explosive having a smaller energy density than a TNT-slurry. The plan was to carry out experiments with both 102-mm and 38-mm hole diameters. The 38-mm hole diameter was to be used to estimate the effect of the rock structure on the breakage. Unfortunately, all of the &##8709;102-mm holes that had been drilled could not be used because of potential interference with the production demands in the mine. Therefore, only two shots could be made in 102-mm holes. Of the 38-mm diameter shots, six holes yielded questionable results. The results reveal that the rock structure plays a major role in the breakage process. Because of the small number of test shots carried out, no evaluation of the proper scaling for production blasts could be made. The tests, however, did show that the hole diameter used for the tests should be chosen as
that the hole diameter used for the tests should be chosen as close as possible to the hole diameter used for production blasting. The rock structure obviously influences the breakage process in a dominant way. However, for guidance, the test shots are reported in Table 72.6.
The Production Blasts The Proposal. This part outlines a proposal based on the performed crater tests. As the reader may have noticed, the experiments with ANFO charges were not carried out to such extent that proper scaling could be done with this type of explosive. However, because of its low price, ANFO is and will be a very competitive explosive whenever dry conditions can be achieved in blasting. Therefore, ANFO should be tested in the Luossavaara mine to make a proper evaluation of the blasting performance.
Although there were not enough ANFO tests for proper scaling, the results achieved, combined with the experience of Leslie Lang, allowed an initial alternative hole pattern relative to the TNT-slurry to be considered. ANFO has lower energy density than a TNT-slurry but produces a larger gas volume per weight of explosive. Because the charge weight only will be around 20 kg, both the energy content and the gas volume release will be much higher with the TNT-slurry. This indicates that the specific drilling will be much higher with the ANFO explosive. Experience from earlier experiments carried out by Lang indicates that the ANFO explosive will need a spacing of 2 m with a loading depth
ANFO explosive will need a spacing of 2 m with a loading depth of 1 m. The crater curve for this type of explosive also indicates a more plastic type of breakage behavior (Figure 72.24). Today, in addition to the high energy TNT-slurries and the ANFO type of explosives, a new generation of explosives exists--the emulsion slurries. Emulsion slurries have been introduced throughout the world and are priced somewhere between the TNT-slurry and the ANFO explosives. The energy density and the gas release energy for this type of explosive are also somewhere between the explosives mentioned. They can be easily loaded up-hole, and they are water-resistant. Today there are three types of explosives, each with a different breaking performance, to study for production blasting--the TNT-slurry, the emulsion slurry and ANFO. If the optimum explosive-geometry-ore combination is to be found, then the following procedure should be followed: 1. Explosive Type 1 (TNT-slurry), Type 2 (Emulsion), and Type 3 (ANFO) are considered for the VCR stope. Explosive Type 1 has a higher density and energy ratio than explosive Type 2, which has higher ratings than explosive Type 3. Consequently do and spacing will be greater for Type 1 than for Type 2 and Type 3. 2. Three different hole patterns are drilled in the stope Dl. Type 1 should be used for the test stope with a hole pattern equal to 3 &##8734; 3 m. Type 2 should be used for a pattern 2.5 &##8734; 2.5 m, and Type 3 should be used for a pattern 2 &##8734; 2 m. 3. The Type 1 explosive should also be tried in the stopes drilled for Type 2 and Type 3 to examine whether the higher specific charge will result in such a good fragmentation that the lower cost for loading, hauling, and crushing will pay for the extra costs caused by the higher specific drilling and specific charge resulting in a lower total mining cost. 4. If the Type 1 explosive pulls satisfactorily in the lower part of the stope with hole pattern 3 &##8734; 3 m, it would be wise to see whether a larger do can be used (a smaller specific charge). The breakage process must be somewhat different when several charges are detonated at the same level and at the same time
detonated at the same level and at the same time compared to the crater experiments where just a single shot is made. 5. To evaluate the results, it is important to follow up the fragmentation distribution achieved and the costs for drilling, blasting, loading, hauling, and crushing. Table 72.7 and Figure 72.25 give the details of the proposal. The suggested values for the Type 2 explosive (emulsion) are uncertain and should probably be checked by a test in 4-in hole diameter with loading depth of 0.6, 0.7, 0.8, and 0.9 m.
Results from the VCR Production Blasts. The main objective of this section was to show a way to measure, perform, and evaluate crater tests. The results from the production stope are summarized in Tables 72.8 and 72.9 to give the reader a feeling of what the outcome became when the test data were applied as design parameters for scaled-up blast.
The explosive used in the production stope had the data found in Table 72.8.
Blast Design for Drifting Introduction Fully automated drifting is not yet a reality, and it will take some time before it is developed to its full potential. Manual work is still needed to charge the rounds. In drilling, scaling, rock support, and mucking, mechanization has improved. In modern mines, hydraulic drilling equipment has taken over after the pneumatic. This has meant enormous capacity increases as the penetration rates have gone up by a factor of 3-4. Most new drill rigs are equipped with air-conditioned and noiseinsulated cabs where the operator may comfortably sit and listen to his choice of music, if so inclined. Computers improve the accuracy and help drill the holes in the face right at the spot where they should be. The earlier, time-consuming survey work and mark-up of the holes to be drilled is, or will soon be, gone. Today, the charging work is not as highly mechanized as the drilling, but there has been considerable development work to increase safety and provide a better work environment. Many mines still use fuse-and-cap or electric detonators, but in modern mines shock-tube detonator systems like Nonel? are a
modern mines shock-tube detonator systems like Nonel? are a must due to safety aspects. With electric detonator systems, there are, unfortunately, still too many accidents due to electrical interference. Many mines perform the drift-charging with scissor-lifts placed on a truck. The charging is done with sticks of NG explosives such as dynamites and cartridges of watergel or emulsion. In dry conditions ANFO is usually charged by means of pneumatic equipment. Formerly, at larger tunnel excavations, one could see drill rigs with one or two booms plus a telescopic working basket that made it possible to charge the holes at the same time as the drilling. In many countries this is considered unsafe and has been abandoned. The risk of falling rock was large and mistakes drilling into a charged hole caused unintentional detonations. Modern mines have invested in dedicated charging equipment, which can visit the face after the drilling is performed, and from where the crew can safely charge the round. The trend in recent years has been away from NG-explosives and toward emulsion explosives. ANFO, however, is still overwhelmingly used. Many emulsion-charging trucks use a repumpable, ready-made emulsion that can be pumped into the boreholes. The most modern of these can bring with them an unsensitized emulsion matrix that will not be sensitized until the emulsion leaves the nozzle of the charging hose. Through chemical gassing, small gas bubbles (hot spots) are introduced that transform the unsensitized matrix to an explosive with excellent characteristics. The advantage of the emulsion is that it is a very good water-resistant explosive. When the detonation is ideal, it produces far less toxic fumes than other explosives, the breakage performance is excellent, and, through smart equipment, the linear charge strength can be made flexible for performing cautious blasting of the contours. The Design Process Dividing the Tunnel Face Area in Design Sections. The basic principles for charge calculations are still based upon the work by Langefors and Kihlström (1963). When the charge calculation of the drill and blast pattern is performed, it is normal to divide the face into five separate sections: • •
Cut section Stoping holes breaking horizontally and upwards
Stoping holes breaking downwards Contour holes Lifter holes The most important operation in the blasting procedure is to create an opening in the face to develop another free surface in the rock. This is the function of the cut holes. If this stage fails, the round can definitely not be considered a success. In the worst case, the rest of the round freezes and cannot be mucked or reshot safely. • • •
Advance. Figure 72.26 shows a cut design in which one largediameter empty hole has been used. The advance is restricted by the diameter of the empty hole and by the hole deviations for the smaller diameter holes. For economy, the entire hole depth must be utilized. Drifting becomes very expensive if the advance (I) is much less than 95% of the drilled hole depth (H).
H = 0.15 + 34.1&##8709; - 39.4&##8709;2 where &##8709; is the large hole diameter expressed in meters (0.05 < &##8709; < 0.25 m) and H is the drilled depth (m). The advance I for 95% advance then is I = 0.95H Equations 72.25 and 72.26 are valid for a drilling deviation not exceeding 2%.
Hole depths needed for a &##8709;102-mm-large hole in a parallel cut would be about 3.2 m. A &##8709;120mm hole would need a hole length of 3.7 m and a &##8709;150 mm will need 4.4 m. For the large hole diameters of &##8709;250 and &##8709;300 mm used in the LKAB tests presented in sections 6.3 and 6.4, the advances were 6.1-6.9 m before scaling for the &##8709;250 mm hole and 7.1-7.5 m for the 300 mm.
Equations 72.25 and 72.26 would predict 5.9 m for the &##8709;250mm and 6.4m for the &##8709;300 mm hole, indicating that the equations underestimate the advances when very large hole diameters are used. The calculations of each individual burden in the quadrangles, the stoping holes, the lifters and the contour holes are described in detail by Persson, Holmberg, and Lee (1994). Number of Blastholes. The number of blastholes necessary to provide a balanced distribution of explosive energy is dependent upon the rock type, the geology, the stress field, the explosive characteristics, blasthole diameter, and the contour blasting requirements. In tunneling and drifting, the number of blastholes and the specific charge used is a function of the drift area. Figure 72.27 gives an indication of the number of boreholes needed for various tunneling areas. Figure 72.28 indicates the required specific energy.
Cut. In the cut, the holes are arranged geometrically in such a way that firing the charges in sequence creates an opening, which becomes wider and wider until the stoping holes can take over. The cut holes can be drilled to form a series of wedges (Vcut), to form a fan (fan cut), or in a parallel-hole geometry they may be drilled in a pattern close to and parallel with an empty, large central hole (parallel-hole-cut or parallel cut). The choice of cut must be made with an eye toward what drilling equipment is available, how narrow the tunnel is, and the desired advance. With V-cuts and fan cuts (where angled holes are drilled), the advance is strictly dependent upon the width of the tunnel. The parallel hole cut with one or two centered largediameter empty holes is being used extensively with large, mechanized drilling rigs.
The advantages are obvious: in narrow tunnels, the large booms cannot be angled sufficiently to create the necessary V-cut angles; it is easier to maintain good directional accuracy in the drilling when all holes are parallel so there is no need to change the angle of the booms. In the parallel cut, standard-diameter holes are drilled with high precision around a larger hole usually with diameters of 65 to 175 mm. The large, empty hole provides a free surface for the smaller holes to work toward, and the opening is enlarged gradually until the stoping holes can take over the breakage. Stoping Holes. After the cut has been shot, the stoping holes will successively enlarge the excavation opening,. The stoping holes have a much easier job to do than the cut holes and the burdens can be increased considerably as the free face to shoot towards is wider. Lifters. When the lifters, the wall holes, and the back holes are drilled, the lookout angle should be considered. For an advance of 4 m, a lookout angle of 3&##176; should be enough for providing space to drill the next round. The floor is seldom dry and as the holes are angled downwards, they are often filled with water. Therefore, a water resistant explosive should be used. It is important to achieve good heave and fragmentation in this part of the round to provide for an acceptable mucking operation. Methods for Minimizing the Damage to the Walls. Optimum results (with respect to cautious blasting through which unwanted damage and smooth perimeter walls are produced) will be achieved when drilling holes are placed at the intended place and when the perimeter holes are shot simultaneously. As indicated, experiments have shown that if adjacent holes are separated in time more than 1 ms, the result deteriorates. Such precise timing will require the use of electronic detonators. Many techniques are used to reduce the linear charge concentration in the contour row and in the buffer row. For example: •
Decoupled plastic pipe charges
Detonating cord String-loaded bulk emulsion Low density/strength bulk explosives (e.g., ANFO or emulsion with Polystyrene) • Notched holes together with a very light charge Successful smooth-blasting requires extremely good precision drilling and a fairly good rock quality. However, it is worth noting that even if the rock mass contains such structural features as bedding planes, joints, and fractures, or if the rock mass contains some poorly consolidated material, a cautious blasting method will always result in less overbreak and less disturbance of the rock mass. Whether the damage affects the stand-up time of the rock contour or not depends upon the character of the damage, the rock structure. the groundwater flow and, last but not least, the orientation of the damaged weakness planes in relation to the contour and the gravity load. • • •
The rock damage can be described by the induced peak particle velocity. This is proportional to the induced rock strain, and it becomes a measure of the damage potential of the wave motion. The surrounding rock mass, of course, contains a number of potential weak planes, each of which is able to withstand a different level of peak particle velocity. It is not unusual for blasters to fail to consider the effects of the charges in the rows adjacent to the often well-planned smoothblasted contour row. Charging the adjacent rows with a heavy charge results in cracks spreading further into the remaining rock than from the smooth-blasted row. It is better to optimize the charge calculations such that the damage zone from the contour holes is limited. This can easily be done by use of Figure 72.29, where the damage zone is given for different linear charge concentrations. Persson, Holmberg, and Lee (1994) provide the equations for calculation of the damage curves.
A burden of 0.8 m is normal for a hole diameter of 48 mm with &##8709;17 mm Gurit pipe charges. From Figure 72.29 it can be seen that this charge results in a damage zone of about 0.3 m. Choosing a fully charged hole of ANFO (charge concentration of l = 1.6 kg/m) in the next row with a damage zone of 1.5 m is of no value because this results in a damage zone that extends 0.4 m further into the rock (1.5 - 0.8 - 0.3 = 0.4 m) than to use a charge concentration that results in a damage zone equal to that caused by the Gurit plus the burden, i.e., 1.2 m. It is apparent from this example that a reduction of the damage zone can be obtained by reducing the charge concentration per meter of drill hole (such a charge should have l = 0.8 - 1.2 kg/m). This obviously results in increased costs for drill and blast operations, but these are balanced, for example, in tunneling by the advantage of a safer roof and decreased costs for grouting and maintenance.
The same exact approach can be used for designing the blasts near stope perimeters. In this case, the curves shown in Figure 72.30 can be used. SveBeFo has performed extensive tests (Olsson and Bergqvist 1993) where they have directly studied the crack lengths from blastholes. Today, several hundreds of boreholes have been blasted in the Vånga granite dimensional stone quarry in southern Sweden. In a 5-m-high bench, three or four identically charged and simultaneously initiated holes were shot. Each hole was primed, the charge length was 4.5 m and the top was unstemmed. Electronic delay detonators (EDS) from Dyno Nobel were used throughout. After the blast, large blocks were carefully removed and cut horizontally using a large diamond circular cutoff saw. The crack pattern was highlighted using a conventional dye penetrant at one or several places along the hole axis. By this direct method SveBeFo has tested the influence of the burden and the spacing, the hole size, the charge concentration, the decoupling ratio, the VOD, and the initiation delay time between holes. The hole diameters used have been mainly &##8709;38, &##8709;51 and &##8709;64 mm. The explosives used were primarily the Swedish contour blasting explosives like Gurit, Kimulux 42, Detonex 80 (80 g/m PETN cord) and Emulet 20 (an emulsion styropore mix with 20% of the volume strength relative to ANFO). The explosives tested have large differences in their VOD values, ranging from about 2,000 m/s for Gurit, to about 4,800 m/s for Kimulux 42, to about 6,500 m/s for Detonex 80. Figure 72.31 gives a comparison of single hole blasting and multi-hole blasting with simultaneous initiation of the holes. Here the basic blasting pattern with a burden (B) and spacing (S) of B &##8734; S = 0.5 &##8734; 0.5 m was used.
Some of the observed results were: Simultaneous initiation with electronic detonators gives much shorter cracks. Thus the cooperation between the different charges has a positive effect on the crack lengths. This is in contrast to the far field vibrations. • The crack length decreases with decreasing coupling ratio. A bulk explosive that completely fills the hole gives the longest cracks. • The crack lengths increase with increasing charge concentration. • When delay times used were as low as 1 ms, the crack lengths still looked more like the cracks from single hole blasts. • Traditional smooth-blasting procedures with conventional LPs give unnecessarily long cracks. Blasthole Sequencing. The charges in a tunnel blast must be initiated in such a sequence that the opening produced by a previous hole can be utilized by the following holes. The initiation sequence is normally as follows: •
1. Cut (in the following order: first quadrangle, second quadrangle, third quadrangle, fourth quadrangle). 2. Stoping (stoping towards the cut and stoping downwards). 3. Contour holes, wall. These holes are shot with the same interval number. 4. Contour holes, roof. These holes are shot with the same interval number. 5. Lifters except corner holes. These holes are shot with the same interval number.
6. Lifter corner holes. Because the rock removed between each hole in the cut and the central empty hole must be blown out to provide expansion room for the rock removed by the next charge, a long enough time interval between these holes is needed for this ejection to occur. The delay times in the cut are usually 75 to 100 ms. For the rest of the round, where the holes have a larger burden, the delay intervals are of the order of 500 ms. Review of Long Rounds Drifting at the LKAB Malmberget Mine Background. Tests with long rounds were conducted in the late 1980s at LKAB in Malmberget (Niklasson, et al. 1988). These tests indicated that it was possible to drill and blast normal drift rounds with a length up to 7.4 m and, therefore, it was recommended that long rounds be introduced, if it could be done with reasonable operating costs. From these tests, the specification for a suitable drill rig was established. This rig would then be tested under operating conditions in the Sofia project during 1990 and 1991 at LKAB Kiruna. The Sofia project has been reported by Niklasson and Keisu (1991). A total length of 1,260 m of drifts were blasted. The project looked at improving the cuts, the rounds themselves, and improving the contour blasting. The Sofia project was divided into two main parts. The first part dealt with short rounds (4.4 m hole depth) and the second part dealt with long rounds (7.8 m hole depth). The standard rounds in Kiruna utilize &##8709;48 mm blastholes and a &##8709;120 mm cut hole (Figure 72.32). Nearly thirty rounds were blasted in part 1 of the project, to serve as a reference for the continuing tests with &##8709;64 mm blastholes.
The second part of the project had the goal of developing a cut so that entire rounds could drilled and blasted with only &##8709;64 mm and no large center hole in the cut. Many types of cuts were tested. As the research work continued, a cut was developed that functioned well for long rounds. Figure 72.33 shows a developed cut providing good advance.
In the Sofia project, Dyno Nobel electronic detonators were used to refine the contour blasting. It was shown that a higher quality of contour was achieved when the contour holes were fired instantaneously. In the contour, the delay time was 5,500 ms with a maximum scatter of 1 ms. Notched holes were also tested in the contour with good results. The notching was performed with a water-jet nozzle. After the Sofia program was completed, the Atlas Copco rig with mechanized rod adding system was introduced into production
mechanized rod adding system was introduced into production in the Malmberget Mine for drilling 7.8 m long holes. The tests by Niklasson and Keisu (1993) in Malmberget were carried out in parallel with production. Drifting was made both in ore and waste rock with considerable variation in rock quality. Of the 220 long rounds drilled and blasted, 115 were monitored in detail. Fjellborg and Olsson (1996) reported additional tests of the long round concept with a large center hole in the cut. The tests conducted at the LKAB Malmberget mine were very encouraging. Parts of this project are described in Section 72.7. Drilling. The drilling pattern in Malmberget was projected on the face with a standard slide projector and manually marked. A portable laser used by the drilling operator gave the reference direction. The drilling pattern was the same as for conventional rounds (short round, &##8709;48 mm). Consequently there was no reduction in the number of holes. The short rounds with &##8709;48 mm holes used a parallel hole cut with a &##8709;102 mm center hole. See Figure 72.32. Malmberget used the cut that was developed at the Sofia project, a &##8709;64 mm opening without a large center hole. See Figure 72.33. Cross bits were used in the ore and button bits in waste rock. Tube steel was used on the middle boom during some weeks of the tests. This gave a stiffer drill string and the accuracy in drilling improved. Charging. As mechanized equipment using a hose feeder had not yet been developed, the conventional charging method was used. Normally in the &##8709;48 mm holes, &##8709;22 mm and &##8709;32 mm pipe charges are used. Pipe charges adjusted to &##8709;64 mm diameter boreholes were not available because this diameter is very unusual in drifting. This resulted in very decoupled charges with a tendency to be easily blown out. A number of undetonated pipe charges were found on the muck pile after blasting. The best result was achieved with the ANFO back blowing technique. This method makes it possible to fill just part of the
technique. This method makes it possible to fill just part of the borehole and thereby reduce the linear charge concentration. Unfortunately this could not be used at all times as ANFO can only be used in dry holes. Kimulux with a diameter of 29 mm was tested in the contour in some of the &##8709;64 mm rounds. The function of the charge was better than when Kimulux 22 mm was used. Less undetonated pipe was found on the muck pile but the blasting seemed to be too powerful. For lifters, when not using ANFO, &##8709;39 mm Dynamex was used. The timing sequence of the cut holes differed from the traditional short rounds. Delays between the holes were much longer to avoid "line up" problems. Scaling. Scaling costs were calculated to be reduced by half when pulling the long rounds. However, the scaling of the face increased considerably, even though the round pulled to full length. Even though the specific charge was much higher for a long round than for a conventional one, no sign of increased damage in the roof or walls was present. These observations were based only on visual observations and on the amount of scaling work that was required. Advance per Round. The authors' report that the performance varied between good to excellent. In summary, 40% of the long rounds functioned very well showing an advance of more than 93%; 40% were fairly good with 90%-93% advance; and remaining rounds were acceptable. The mean advance per round was 7.0 m. Poor advance mainly depended on some factors such as defects in the rock or an abnormal amount of water. Hole deviations were reported to be of no problem when &##8709;64 mm Retrac bits were used. Figure 72.34 shows the advances for long rounds in Malmberget.
Conclusions from the Results of the Malmberget Test. Opening cuts with only &##8709;64 mm holes work well, just as well as standard cuts with large-diameter empty center holes. These tests apply for short rounds with 4.0-4.5 m drilled depths, both in ore and in bedrock. • Long rounds, 7.8-m long, were found to be economically feasible for introduction at LKAB in Malmberget. • By precisely delayed intervals or radial notching, the quality of the contour could be improved considerably. • The contour test showed that standard explosive products suitable for the &##8709;64 mm holes do not yet exist. This is to be expected because this diameter is not normally used, and this might be the reason why the contours of the 48 mm rounds generally were of better quality than those with &##8709;64 mm holes. • A precise laser reference for the alignment instruments and an accurate marking of the drilling pattern on the face are very important factors, not only to keep the overbreak low but also to keep the drift heading in the right direction at the right level. • A clean floor in front of the face is required to obtain a correct lookout angle for the lifter holes. •
Substitution of the &##8709;64 mm Cut by a &##8709;300 mm Center Hole Background. Introducing very large-scale sublevel caving has led to higher productivity and lower costs for LKAB's mines in Kiruna and Malmberget.
Kiruna and Malmberget. Development is still the most expensive unit operation. Large scale means that the number of available development faces on each level is very limited, which means that the requirement for effective drift driving is pronounced. In the work cycle for drifting, up to ten different activities are needed. The benefit of using long rounds is obvious as the total set-up time for pulling twice the conventional advances is reduced by 50%. To increase the effectiveness, the length of the round has been increased as mentioned in section 72.6.3. In 1993 trials were conducted with various cut geometries based upon the use of &##8709;64 mm holes. One of the cuts (Figure 72.35) provided advances up to 95% of the drilled length. This cut had six uncharged &##8709;64 mm holes to provide for a better swell volume. The geometry of the cut also meant that the distance between charged holes increased and thereby the risk for dynamic dead-pressing of the explosive was reduced.
Cut performance tests clearly showed that when a &##8709;300 mm empty hole was introduced in the cut the advance per round increased to 100%. The cost for the largediameter central hole could be balanced by the cost savings associated with less drilling, explosive use, face scaling, and a better quality of the contour. In 1994, LKAB decided to purchase a special drill rig for drilling this large-diameter cut hole. A project was established around the use of this drilling machine to:
the use of this drilling machine to: Optimize the diameter of the large diameter hole Find the best blasting plan Determine the blast damage zone Reduce scaling and reinforcement Project Goal. The main goal of the project was to use a predrilled large-diameter cut hole to optimize the length of the rounds, refine the work-cycle, maximize the advance, and minimize the blast damage zone. The target was to achieve 99% advance in 90% of the rounds. Drilling error for the contour holes should be limited to 20 cm outside the planned contour, and the scaling and reinforcement costs should be reduced by 30%. • • • •
Test Area. The main testing area was in the Norra Alliansen orebody on the 790-m level. Four 150-m-long drifts mainly in the ore were allocated for the tests. The magnetite orebody dipped at 45&##176;. The footwall waste rock was a leptite and a low-strength biotite schist with layers having a thickness of a few centimeters up to 1.5 meters. Drilling Equipment. The drilling rig purchased for the largediameter cut hole was an AMV equipped with a 6-in Wassara ITH water-powered machine (Figure 72.36). The large-diameter hole was drilled in two steps. A &##8709;165 mm pilot hole was drilled first and then reamed to a diameter of 250 or 300 mm. The maximum hole length was determined to be 32 m based upon an estimated maximum hole deviation of 1%. The tube magazine contained 25 pieces of 2-m-long drill tubes. The drill penetration rate for the &##8709;165-mm-diameter pilot hole in magnetite was 0.3-0.4 meter/min. The drill penetration rate when reaming the pilot hole to full size was 0.17 meter/min for the &##8709;250 mm and 0.11 meter/min for the &##8709;300 mm large hole. An Atlas Copco Rocket Boomer 353S (Figure 72.37) equipped with an automatic rod adding system (RAS) and Bever Control tunneling position system were used to drill the long drift rounds. The average depth of drilled boreholes was 7.5 m. Part 1--Optimal Diameter of the Large Hole. The standard drilling pattern for the &##8709;64 mm holes was used (see Figure 72.38). The standard drift round has a size of 6.5
Figure 72.38). The standard drift round has a size of 6.5 &##8734; 5.0 m and contains 57 holes. The predicted advance is 7.5 m. All holes except the back holes are charged with the non-cap-sensitive pumpable water-resistant emulsion explosive Kimulux R and a KP primer (VOD 7500 m/s). The back holes are charged with a small 0.5 m long bottom charge of emulsion plus a 40 g/m detonating cord KSP40.
The objective of this part of the project was to compare the advance between the standard long drift rounds and rounds containing a large center hole of &##8709;250 mm or &##8709;300 mm. When testing the large-diameter hole, the center of the standard cut was replaced by the large hole
center of the standard cut was replaced by the large hole (Figure 72.39).
These tests included 14 standard rounds, 7 rounds with the &##8709;250-mm-diameter hole and 5 rounds with the &##8709;300 mm hole. Part 2--Optimal Blasting Plan. This part of the project included tests with a drill pattern where every borehole is located based upon the expected rock removal produced by that hole. Each hole should have its own optimal burden. The hole positions and the delays around the cut are set so the blasting sequence becomes "corkscrew-shaped" (see Figure 72.40).
Tests were made using electronic detonators (EDS), 40 or 80g/m detonating cord and string-loaded Kimulux R in the contour holes (Table 72.10). The tests with electronic detonators included tests of the two systems developed by DNAG and Dyno Nobel.
Nobel. The Associated Blast Damage Zone. Both borehole logging and the slot technique were performed to determine the amount of damage induced in the contour rock (Table 72.11). Borehole logging was applied in four diamond-drill holes and cracks were observed before and after drifting by a borehole logging TV camera. The tests using the slotting technique were performed by SveBeFo who have applied this technique (Olsson and Bergqvist 1996) over many years to study blast damage.
After blasting, a special diamond saw was used to make a series of vertical cuts 2 m long and 0.5 m deep.The rock between the cuts was then removed and the rock surface perpendicular to the bore hole axis could be examined for radial blast-induced cracks. A dye penetrant was sprayed on the surface and photos were taken of the crack patterns. Results. The advances were measured before and after scaling and compared with the standard reference 65-mm long drift rounds. Scaling was performed with a Montabert BRP 30 hydraulic hammer and with water at 100 bars pressure. The results clearly showed that using a large-diameter cut hole increased the advance when using the standard drill pattern (Figure 72.41). The advance for the &##8709;300 mm holes, even before scaling, had an average advance of 97%. It should also be noted that the rounds with large-hole-diameter central holes have a better advance before scaling compared to the standard reference round after scaling.
In all of the blasted rounds, hole deviation was measured in about 20 bore holes per round (Table 72.12). The deviation varied from zero to a maximum deviation of 0.4 m. The deviation is equally spread over the face but with the largest deviation observed in the contour holes. The average is 2.8% for the standard rounds while that for the large center cut hole rounds is 2.2%. The appearance of half casts varies greatly with the type of explosive, the ignition system, and their different combinations (Figure 72.42). Contour holes fully charged with emulsion and contour holes charged with 80g/m detonating cord when initiated by long period delay (LP) caps showed almost the same result of half casts. Standard LP caps with a delay of 3,500 ms were used in the whole contour. Detonating cord used in combination with electronic caps (EDS) gave results that were significantly better due to the instantaneous ignition. The best overall results based upon half cast observations were obtained when using electronic caps (EDS).
The method with the slot combined with dye penetrants was successful in examining blast damage. The results could very clearly distinguish crack patterns with the different explosives and initiation combinations. Contour blasting with a decoupled string of emulsion initiated instantaneously with electronic detonators resulted in no blast initiated cracks. Contour holes fully charged with emulsion ended up with radial crack lengths of at least 0.5 m. The results of the cautious contour blasting tests indicate that, when electronic detonators are used, the type of explosive used has a minor influence on the results. The excellent results depend mainly upon the instantaneous ignition with scatter below one millisecond. The amount of overbreak was measured for every round. The amount of overbreak depends on a number of factors such as the alignment of the drilling rig, the drilling accuracy, the method of scaling, and the geology. Therefore, it cannot be used as a true measure of the blasting quality, but it does gives a good indication of how the overall operation is developing. The average overbreak at the beginning of the project was greater than 15%, sometimes reaching 30%. At the end of the project, overbreak had decreased to an average of 12%. This project, during the period of September 1994 through November 1995, included a total of 53 drift rounds involving the use of a large-diameter hole. By achieving an average advance of 99.5% and by cutting the need for scaling by 50% when using a large-diameter cut hole in combination with the modified drilling pattern, the main goals of the project were fulfilled. Based upon the project results, LKAB has introduced the predrilled large-diameter cut hole method in the Malmberget mine.
drilled large-diameter cut hole method in the Malmberget mine. Some Environmental Issues Good Blasting Practice Is Important. To minimize any spillage during the charging and blasting, it is essential to be aware of some factors that may affect the aquatic environment. Operators and management should pay attention to the following: A good drilling accuracy is important. Too short distances between the holes can result in dynamic dead pressing of the adjacent hole causing undetonated explosives in the muck pile. • Do not charge nonwaterproof explosives such as ANFO into wet or water-filled boreholes. • Bulk explosive may be discharged when one carelessly moves loading hoses from hole to hole. • When ANFO is being loaded pneumatically, avoid any blowback during the charging or any spillage when the hose is transferred to the next hole. • Be aware of excess ANFO at the collar that will get into the muck pile. • Spillage may occur when charging or recharging loading equipment on site or cleaning the equipment after performing the operation. • Do not empty loading hoses by flushing the explosive onto the ground. • Proper timing is essential as cutoff boreholes can result in undetected explosives in the muck pile presenting a potential contamination hazard. Long exposure to water and warm temperatures increases the rate of the reaction from ammonium nitrate to ammonia and nitrite. Therefore, it is of utmost importance to test the quality of water on blasting sites. Water of doubtful quality should never be released into natural water systems. •
All civil explosives contain ammonium nitrate. Products originating from chemical reaction of ammonium nitrate may present a risk to the aquatic environment if nondetonated explosive comes into contact with water and the ammonium nitrate dissolves. The dissolved ammonium nitrate may be transformed into nitrite or ammonia. At mines, quarries, and tunneling work usually there are limits to what the owner/contractor can release to the environment and it is
owner/contractor can release to the environment and it is always a good practice to pay attention to the drilling and blasting practice and the explosives used in order to minimize any contamination. Depending on their nitroglycerine content, gelatinous explosives have good-to-excellent water resistance. The water resistance of emulsion and watergel explosives is excellent. In addition, extra protection can be obtained from the cartridge wrapping. The higher the degree of water resistance of an explosive, the lower is the risk of contamination. Spillage is not a problem when using cartridged explosive. Powder explosives containing nitroglycerine or TNT form the majority of nonwaterproof cartridge explosives. The explosives' compositions themselves are not inherently waterproof although the cartridge wrapping affords some protection but, in general, cartridging only affords a small measure of protection, since the cartridge may be damaged during the loading operation Waterproof bulk explosives include watergels and emulsions, and these may resist water for several weeks or even months. Spilled emulsion or watergel explosives will dissolve slowly. Dissolving will be faster after the emulsion has been exposed to mechanical stresses. In the case of emulsion explosive, such stresses may break down the emulsion and separate salts such as ammonium nitrate from the oil and water. Subsequently the salts may dissolve in water. The main nonwaterproof bulk explosive in this category is ANFO, which will dissolve easily in water and should not be used in wet or water-filled holes. Toxic Fumes. Fumes are the gases resulting from detonation. Typically, one kilogram of explosives will produce between 700 to 1,000 liters of such gases. The stable products of detonation are nitrogen, carbon dioxide, and water, but in addition small quantities of carbon monoxide and nitrous gases are produced. Toxic fumes amounts to around 4% of the after-detonation gases. CO typically &##126;3% and NOx &##126;1%. The amount of nonideal detonation products formed depends on a number of factors: the type of explosive, the water resistance, type of cartridge wrapping, the VOD, the charge diameter,
type of cartridge wrapping, the VOD, the charge diameter, loading density, type of initiation, and especially the confinement of the explosives. Some recommendations to minimize the toxic fumes: Be careful to drill the holes at the right position according to the drilling plan. This gives less toxic fumes and best blasting results. • Use alignment devices when drilling so most holes are slanted slightly upwards. This prevents water from accumulating in the drill hole, contaminating the explosives and affecting the detonation properties. • Use an oxygen-balanced explosive with good fume characteristics. • Leave an unloaded hole length or stem the holes! • Explosives in the collar increase the amount of toxic fumes--but not the breakage! • Avoid cord in ANFO as it might not initiate ANFO to complete reaction, resulting in toxic fumes. The cord itself is strongly oxygen deficient, and by itself generates about 3-l CO per meter cord. • Explosives in "air" increase the fumes. • Avoid spacers between cartridges! • Considerable quantities of after-detonation fumes can become trapped in the muckpile. A good practice is to flush the muckpile with water to remove the dust and the trapped gases before mucking and hauling start. • Good shot-firing practice contributes towards balanced fumes from blasting operations but does not remove the need for proper and adequate ventilation. • Measure the toxic fume concentration before entering the mine after a blast •
ACKNOWLEDGEMENTS The authors wish to thank the management of AECI Explosives & Chemicals Limited for permission to include the section on ring blasting.
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1. Dyno Nobel, Gyttorp, Nora, Sweden. 2. Dept. of Mining Engineering, University of Utah, Salt Lake City, Utah. 3. AECI Explosives Ltd., Modderfontein, South Africa.
