Empirical Equations Supporting Blast Designs Discussion on Design Parameters and Related Matters

EMPIRICAL EQUATIONS SUPPORTING BLAST DESIGNS Discussion on Design Parameters and Related Matters *** Partha Das Sharma, B.Tech(Hons.) in Mining Engineering; E.mail: [email protected]; Weblog: http://miningandblasting.wordpress.com/ INTRODUCTION: Fragmentation is an important factor to qualify a blast, because it has a clear impact upon profit and productivity of a mining operation. Apart from technical results, related to safety and environment such as vibration and projection, results are essentially considered on an economical point of view. Thus, there are other factors as well which impact upon profit and productivity. Majority of all these factors dependent on how well a blast is designed. In other words, performance of a blast mostly depends on Blast design. The primary objectives in rock blasting are to optimize blast performance and ensure enough safety in operation. With the continued evolution of drilling equipment, and the extension of surface mining, bench blasting is fast becoming the most popular method of rock fragmentation with explosives. Many formulas and methods for calculating geometric parameters such as burden, spacing, and sub drilling have been around since the early 1950’s. In small diameter blasting the most common technique developed by Langefors and Kihlstrom is used; however, it is better to use the crater technique by Livingston or the American criteria for the larger diameter blasts. Due to the different nature of rocks the best method is continuous trial and error to arrive at the best conclusion. Obviously, every situation in the field cannot be predicted, and is beyond the scope of this chapter. What this chapter will do is give an initial approach to the approximate geometric design of blasting, the calculation of charges, and characterization of rocks by their uniaxial compressive strengths (UCS). It will be necessary to adjust patterns, explosive charges to suit the need in the field according to the type and makeup of the material encountered. EMPIRICAL EQUATIONS SUPPORTING BLAST DESIGNS: * Fraenkel (1944)

Where, B max = Maximum burden for good fragmentation, m d = Borehole diameter, m

EMPIRICAL EQUATIONS SUPPORTING BLAST DESIGNS Discussion on Design Parameters and Related Matters

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hc = Charge height, m H = Depth of the blast hole, m * Andersen (1952) determined the burden value in feet and its value increases with the length of the blast hole but not indefinitely as usual happens in practice.

Where, B = Burden, ft D’ = Diameter of hole, ft L = Length of the blast hole, ft K = Empirical constant This formula does not take into account the rock properties or those of the explosives. * Pearse (1955)

B = maximum e burden (m) K = Constant, value varies from 0.7-1.0 Ps = Detonation pressure of the explosives (Kg/cm2) σt = Tensile strength (Kg/cm2) d = Diameter of borehole * Hino (1959) The equation proposed by Hino is:

B = Burden, m D = Blasthole diameter, cm PD = Detonating Pressure, Kg/cm2 RT’ = Dynamic Tensile Strength, Kg/cm2

Partha Das Sharma, B.Tech(Hons.) in Mining Engineering; E.mail: [email protected]; Weblog: http://miningandblasting.wordpress.com/


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N = Characteristics constant depending upon the par explosive-rock and calculated through the catering test.

Where, D’ = Optimum depth of the center of gravity of the charge, cm and it determined graphically from the following equation values,

D = diameter of the explosive charge D” = depth of the center of gravity of charge Δ = Relationship of depths D”/Dc Dc = Critical depth of the center of gravity of charge Σ = Volumetric constant of charge V’ = Volume of charge used * Allsman (1960) The equation for maximum burden value proposed is;

Where, PD = Mean adverse detonating Pressure, N/m2 T = Duration of average detonation, sec Ρ = Specific rock weight, N/m3 U = minimum velocity which must be imparted to the rock, m/s G = acceleration due to gravity=9.81 m/s2 D = Diameter of blasthole, m * Ash (1963) Burden, B (ft) = 0.084 × KB × D (in)



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Where, KB = Depends upon the rock group and the type of explosive used, Blast hole depth, L= KL× B (KL between 1.5 & 4) Sub drilling, J = KJ× B (KJ between 0.2 & 0.4) Stemming, T = KT× B (KT between 0.7 & 1) Spacing, S=Ks× B Ks = 2.0 for simultaneous initiation, 1.0 for sequenced blasthole with long delay between 1.2 & 1.8 for sequenced blasthole with Short delay * Langefors and Kihlstrom (1968)

Where, B max = Maximum burden for good fragmentation (m) D = diameter of hole (m) ρe = Density of the explosive in the borehole (Kg/m3) PRP = Relative Weight strength of the explosive f = Degree of confinement of the blasthole. S/B = Spacing to burden ratio Co = Corrected blastability factor (Kg/m3) = C + 0.75 for B max =l.4-1.5m = C + 0.07/B for B max < 1.4m When C = rock constant * Lopez Jimeno, E (1980) He modifies the ash’s formula by incorporating the seismic velocity to the rock mass, resulting in B=0.76XDXF Where, B = Burden, m D = Diameter of blasthole, inches



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F = Correction factor based on rock group = Fr× Fe

Where, ρ' = specific gravity of rock, gm/cm3 VC = seismic propagation velocity of the rock mass ρ'' = specific gravity of explosive charge, gm/cm3 VD = Detonation velocity of explosive, m/s The indicated formula is valid for diameter between 165 & 250mm.For large blasthole the burden value will be affected by a reducing coefficient of 0.9. * Konya and Walter (1990)

Where, B = Burden, (ft) ρe = Specific gravity of explosive, (lb/in3) ρr = Specific gravity of rock, (lb/in3) D = Diameter of explosive, (in) Correction factor, Bc = Kd. Ks. Kr. B Where, Bc = Corrected burden (ft) Kd = Correction factor for rock deposition. Its value is as follows, • for bedding steeply dipping into cut Kd = 1. 18



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• for bedding steeply dipping into face Kd = 0.95 • for other cases Kd = 1.0 Ks = Correction factor for geologic structure. Its value is as follows, • for heavily cracked, frequent weak joints, weakly cemented layers Ks = 1.30 • for thin well cemented layers with tight joints Ks=1.1 • for massive intact rock Ks = 0.95 Kr = Correction factors for number of row. Its value is a follows, • for one or two rows of blastholes Kr = 1.0 • for third or subsequent rows Kr = 0.95 Konya and walter also suggested the following empirical relationshipsFor instantaneous initiations system,

For delay initiation system,

Where, H = depth of blast-hole, m B = burden, m S = Spacing, m Konya and Walter also suggested the following empirical relationship-

Where, SANFO = relative strength of explosive ρr = density of rock, gm/c.c.


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EMPIRICAL EQUATIONS SUPPORTING BLAST DESIGNS Discussion on Design Parameters and Related Matters d = diameter of blast-hole, m Drillability and Blastability are the geological influences in hard rock drill and blast: The drillability is not only decisive for the wear of tools and equipment but is – along with the drilling velocity – a standard factor for the progress of excavation works and blasting performance. The estimation of drillability in predicted rock conditions might bear an extensive risk of costs. Therefore, an improved prediction of drilling velocity and bit wear would be desireable. The drillability of a rock mass is determined by various geological and mechanical parameters. Drilling velocity is dependent on a lot of geological parameters: Those principal parameters include jointing of rock mass, orientation of schistosity (rock anisotropy), degree of interlocking of microstructures, porosity and quality of cementation of clastic rock, degree of hydrothermal decomposition and weathering of a rock mass. Drilling bit wear increases with the equivalent quartz content. The equivalent quartz content builds the main property for the content of wear-relevant minerals. For various groups of rock types different connections with the equivalent quartz content could be detected. In sandstone bit wear is also dependent on porosity or the quality of the cementation. Rock Blastability: The term “blastability” is used to indicate the susceptibility of the rock mass to blasting and is closely related with the powder factor. An attempt to relate blastability to rock and rockmass properties has been reported Lilly (1986). The blastability index (BI) is defined as:

Where, RMD = Rockmass description, JPS = Joint plane spacing, JPO = Joint plane orientation, SGI = Sp. Gravity influence and H = Hardness (Moh’s scale of hardness). Rating of different parameters is given in following tables. ANFO powder factor and Energy factor consideration has been given with BI in the graph (as per field data).

Note: A BI value of 100 refers to a massive, extremely hard, iron-rick cap rock. It has a Sp. gravity of 4. Soft, friable shale has an index around 20.


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Ref: Lilly,P.a., (1986), An empirical method of assessing rock mass blastability, Proc. Large open pit mining conf. (J.R.Davidson ed.), The Aus. IMM, Parkville, Victoria, October pp. 89-92.

DISCUSSION ON BLAST DESIGN: The important parameters which generally govern for blast designs are: * Physico-mechanical properties of rock: Here type of the rock, dynamic tensile strength, tensile strength, compressive strength, young’s modulus, Poisson’s ratio, density and hardness of the rock mass, presence of discontinuities, bedding plane and joints, etc. are very important. * Pit geometry: Factors considered are thickness of coal seam or ore body and bench height, over burden bench height, bench slope angle, strip width, height to width ratio, and length to width ratio are generally considered. * Explosive characteristics: Factors generally considered are type of explosive, type of booster, bulk strength, energy release per unit mass of explosive, detonation pressure, explosion pressure, ratio of decoupling, strength of explosive used, time taken for explosive wave to travel to the free face and back, volume of gaseous product per unit mass of the explosive, velocity of detonation, velocity of explosion propagation, explosion wave length, weight strength, number of spalls that an explosive wave may produce, length, diameter and weight of the cartridge, loading density, bottom charge and column charge density, etc. are very important. Characteristics of blasting accessories - type, thermal properties are also important. * Blasting technique: Here objective of blasting, drilling pattern, number of availability of free faces, manner of charging, charge per hole and per delay, sequencing of initiation i.e. delay between two holes in a row and delay between two rows, decking, length of explosive column, height of the bottom charge, volume of the explosive in the blast hole, etc., are to be considered.



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Important Parameters to be considered: · Burden distance. · Spacing of the hole. · Ratio of spacing to burden. · Depth of hole. · Diameter of blast holes. · Consideration of toe and depth of sub-grade drilling.

* Powder factor: The size of the fragmented rock should match the bucket size of the excavator and also the grizzly size of the primary crusher. Important Parameters to be looked into: · Length of stemming column, the size and quality of stemming · Angle drilling · Amount and direction of throw requirement and problems of fly rock. · Requirement of muck profile · Vibration level · Presence of water

BENCH GEOMETRY * Bench Height (H): The bench height is the vertical distance between each horizontal level of the pit. Unless geologic conditions dictate otherwise, all benches should have the same height. The height will depend on the physical characteristics of the deposit; the degree of selectivity required in separating the ore and waste with the loading equipment; the rate of production; the size and type of equipment to meet the production requirements; and the climatic conditions. The elements of a bench are illustrated in the figure below. The bench height should be set as high as possible within the limits of the size and type of equipment selected for the desired production.



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The bench should not be so high that it will present safety problems of towering banks of blasted or unblasted material or of frost slabs in winter. The bench height in open pit mines will normally range from 15 m in large copper mines to as little as 1 m in uranium mines. But in special case such as rip-rap blasting height can be reached 20 m. The bench height is directly related to degree of heaping and spreading of material broken by blasting, thus, directly affecting displacement requirement to accomplished by round blasting. The height also limits the maximum and minimum charge diameters and drill diameters. The most economical may be also determined by the drill penetration rate; whenever penetration rate decreases significantly, it is generally uneconomical to drill deeper. High faces pose the problem of considerable bit wander, especially with small diameter hole. The deviation of blast hole places a limit on the maximum allowable bench height. The bench height is also highly depend on capacity of loading equipment. The following are some of the factor that should be considered in the selection of the bench height: Optimum blast hole diameter increases with the height. In general an increase in blast hole diameter decreases in drilling costs. In some cases the bench height is limited by the geology of the ore deposit due to imperatives of the ore dilution of the control and safety measures. There are a couple of ways to calculate the bench height of a large diameter blast hole, the first of which relates to the size and reach of the rope shovel. The height in meters can be estimated by the following equation: H = 10 + 0.57 (Cc– 6) Where, Cc = the bucket size of the shovel (m3) and H = bench height (m) Another way to calculate bench height which take into account the compressive rock strength and relate it to the diameter can be seen in Table:



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Relationship of bench height, stemming with diameter & UCS of rock Design Parameter Compressive rock strength (MPa) Low Medium-high Very High < 70 70-180 >180 Bench Height H 52 x D 44 x D 37 x D Stemming - T 40 x D 32 x D 25 x D

* Bench Width: There is a minimum bench width, measured horizontally in a direction perpendicular to the pit wall. For each bench height and set of pit operation conations whose value is established by the working requirements of the loading and hauling equipments. The width also must be such so that to ensure stability of excavation both before and after blasting, because each blast effectively reduced the restraint sustains the pit walls at higher elevation. Because of the limit set by requirements for equipments operating room and bank stability, there is a maximum width that should not be exceeded by any blast. BLAST GEOMETRY * Drilling Diameter (D): The hole diameter is selected such that in combination with appropriate positioning of the holes, will give proper fragmentation suitable for loading, transportation equipment and crusher used. Additional factor that should be considered in the determination of the hole diameter are: a) Bench height, b) Type of explosive, c) Rock characteristics, d) Average production per hour.

The drilling and blasting will become economical with increase in diameter. When the blast hole diameter is increased & the powder factor remains constant the large blast hole pattern gives coarser fragmentation. By keeping burden unchanged & elongating spacing alone the problem can be overcome. When joins or bedding plane divide the burned into larger blocks or hard boulder lie in a matrix of softer strata acceptable fragmentation is achieved only when each boulders has a blast hole, which necessitates the use of small diameter blast holes. Hole diameter varies from 35 in small benches up to 440 mm in large benches. In India 100-150 mm blast hole diameter are used in limestone mines,150-270 mm in coal mines & 160 mm or above blast hole are used in iron ore mines is used. Langefors and Kihlstrom suggested that the diameter be kept between 0.5 to 1.25 percent of the bench height.


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EMPIRICAL EQUATIONS SUPPORTING BLAST DESIGNS Discussion on Design Parameters and Related Matters Diameters from 165 mm to 450 mm are considered to be large diameter bench blasts. Large diameter bench blasts are used mostly in large surface mining operations and certain civil engineering excavations like power stations and quarries for the construction of dams. Many of the same variables are required for the proper calculations. The average production per hour and type of rock being fragmented is still the variables needed for consideration. * Sub Drilling (J): Sub drilling is the length of the blasthole underneath the floor level needed to break the rock at bench height and achieve adequate fragmentation and displacement; this allows the loading equipment to achieve optimum level of productivity. However, sub drilling is not used in calculating the volume of rock being blasted. If sub drilling is too small, the rock will not completely shear off resulting in a toe appearance (this leads to an increase in loading costs). However, if the sub drilling is too large the following can happen: a) Increase in drilling and blasting costs, b) An increased vibration level, c) Excessive fragmentation of the bench, affecting slope stability in the end zones, d) Increased risk of cutoffs and over break. The optimum effective sub drilling depends on a) The structural formation, b) Density of the rock, c) Type of explosive, d) Blasthole diameter & inclination, e) Effective burden. Sub drilling is usually calculated from blasthole diameter: J=5+ (0.450-D)/0.09467. In general, the value of sub drilling that produces the optimum level of breakage is roughly 0.3 times of Burden. When drilling vertical blasthole the first row should reach values of approximately 10 to 12 times D. Shorter lengths then about 7 – 8 x D for 180 to 250 mm blast hole, and 5 – 6 x D for 250 to 450mm blasthole or less, if used in the cases: a)Horizontal bedding planes that coincide with the bench toe, b) Application of select explosive charges. * Location of initiators in the charge: It is usually calculated from blast hole diameter when vertical blast holes are drilled. The sub drilling of the first row reaches value of 10D to 12D. About 10% of sub drilling gives better fragmentation in the rock mass and lesser ground vibration. In generally sub drilling should be 0.3 times the burden. Under different toe conditions sub drilling may be up to 50 percent of the burden. A relation is also shown in the figure below.


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EMPIRICAL EQUATIONS SUPPORTING BLAST DESIGNS Discussion on Design Parameters and Related Matters Excessive sub grade drilling causes more vibrations, under fracturing at the bottom and depressed floor conditions. It should be avoided since; a) It waists drilling and explosives expenditure, b) Increased ground vibration level may cause undesirable shattering of the pit floor, c) Increase the vertical movement of the blast. * Stemming (T): The primary function of the stemming is to confine the gas produced by the explosive until they have adequate time to fracture and move the ground. A suitable stemming column of suitable length and consistency enhances fracture & displacement by gas energy. The amount of unloaded collar required for stemming is generally from one half to two third of the burden, this length of stemming usually maintains sufficient control over the generation of the objectionable air blast, fly rock from the collar zone. When the burden has a high frequency of natural crack and planes of weakness relatively long stemming column can be used. When the rock is hard and massive the stemming should be shortest which will prevent excessive noise, air blast and back brake. For blast hole diameter in the 230-380 mm range, angular crushed rock in the approximate size of 23 to 30 makes a very effective stemming column. Larger fragments tends to damage the detonating cord and the detonator lead wire. In coal blast inert stemming material should be used rather than coal cutting. In multi row blast when the mean direction of rock movement tends to more and more towards the vertical with successive rows a longer stemming column is often used in the last row to avoid over break. When large stemming is kept in rocks with discontinuities large boulders may result. In such cases pocket charge or satellite charge are recommended. Just as with any other calculations, this (stemming length) too must be accurate. If the stemming is too great (excessive) then this leads to a large quantity of boulders coming from the top of the bench, poor swelling of the muck pile, and an elevated vibration level. However, if the stemming is too small (insufficient) then this leads to a premature escape of the gases leading to an air blast and a danger of fly rock, the hurling of rock fragments in a blast. To properly calculate stemming, the type and size of material used, and the length of the stemming column must be taken into account. Studies have shown that coarse angular material, such as crushed rock, is the most effective stemming product. Crushed rock effectively lowers the stemming length by up to 41%. The optimal stemming length varies between 20 and 60 times the diameter of the blast hole with at least 25 times the diameter maintained to avoid the problems. From the field experience, it is realized that stemming length of 70 percent of the burden dimension a good approximation. This length has a sufficient control over production of objectionable air blast and fly rock from the Collar zone. It is recommended that the crushed and sized angular rock fragments works best as stemming. But it is common practice to use drill cuttings as a stemming material. * Blast Hole Inclination (β): In recent year attention has been given by open pit operators to the drilling of blast holes up to 20 degree vertical. The benefits from inclined charges are a) Reduction of collar and toe region, b) Less sub drilling requirement, c) Uniformity of burden throughout the length of blast hole, d) Drilling of next bench is easier. Air blast and fire rock may occur more easily due to smaller volume of material surrounding the collar inclined hole are successively used in Europe where high benches and smaller diameter holes in medium to higher strength rock exist. In case the face is high the use of vertical blast holes produce a considerable variation in burden between the top and bottom face which is the basic cause in the formation of toe.


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EMPIRICAL EQUATIONS SUPPORTING BLAST DESIGNS Discussion on Design Parameters and Related Matters Angle greater than 25 degree are less used because of difficulty in maintaining blast hole alignment excessive bit wear and difficulty in charging blast holes. The blast hole length L increases with inclination. To calculate L, the following equation is used:

Where, β in degrees represents the angle with respect to the vertical. In bench blasting it has been discovered that inclined drilling gives the most benefits with few disadvantages. Some of the benefits include: a) better fragmentation, b) less sub drilling, c) increased drilling productivity, and d) lower powder factor. Some of the disadvantages are a) an increased drilling length, b) more wear on bits, and c) problems in charging the explosive. The blasthole length increases with inclination; however, the sub drilling decreases. Burden (B): In surface mines, a successful blast design provides a sufficient quantity of suitably fragmented and properly placed muck at the lowest practical cost. One of the significant variables in such blast designs is the 'burden'. This is one of the most critical parameter in designing of blast. It is the distance from a charge axis to the nearest free face at the time of detonation. As the boreholes with lower delay periods detonates, they create new free faces. As a result the effective burden will depend upon the selection of the delay pattern. When the distance between discontinuities is larger, smaller burden is required. A relationship between burden with blast hole diameter has been shown in the figure below.

* Spacing (S): Spacing is an important parameter in blast design. It is defined as the distance between any two adjacent charges in the same row and it controls mutual stress effect between charges. Spacing is calculated as a function of burden, hole depth, relative primer location between adjacent charges and depends upon initiation time interval. Over past several decades in most mining operations the spacing distances have been decided in relation to burden. The value of the spacing to burden ratio (S: B) which has been commonly used in different formulas lies between 1 and 2. From the production scale test with



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the spherical charges breaking to crater geometry, many workers suggested that the spacing be kept about 1.3 times the burden. When this ratio increases more than 2, unexpected results were found. These (Burden and Spacing) parameters are dependent on the following variables: drilling diameter, properties of the rock and explosive, the height of the bench, and the degree of fragmentation and displacement. There are many formulas that have been suggested for calculating the burden, taking into accounts one or more of the variables mentioned. If the burden is too great, then the explosion gases encounter too much resistance to effectively fracture and displace the rock. Part of the energy used is turned into seismic energy and intensifies ground vibration. This is most evident in pre-splitting blasts where there is total confinement and vibration levels can be as much as 5 times larger then normal bench blasting. If the burden is not large enough, the gases escape and expand at high speeds towards the free face. This pushes the fragmented rock, and projects it uncontrollably causing an increase in overpressure of the air and noise. The spacing S value is calculated with burden and the delay timing between blasthole. The value for spacing is approximately 1.15 x B for hard rocks, and 1.30 x B for soft rocks. As with burden, if the dimensions for spacing are inadequate then irregularities occur in the rock face. If the spacing is too large then the fracturing between the charges is inadequate and leads to toe problems. If the spacing is too close together then excessive crushing between charges occurs, along with superficial crater breakage, large blocks in front of the blast hole, and toe problems. * Drill Patterns: The burden as indicated previously is a function with the charge diameter, compressive rock strength, and specific energy of the explosive used. The diameter of the column charge is usually the same as the drilling diameter. List of burden and spacing values for various compressive rock strengths and explosives are given in Table: Burden and spacing values for various compressive rock strengths and explosives Design Compressive rock strength (MPa) Parameter Soft Medium-Hard Very Hard < 70 70-180 > 180 ANFO Burden - B 28 x D 23 x D 21 x D Spacing - S 33 x D 27 x D 24 x D Water gel / Burden - B 38 x D 32 x D 30 x D Emulsion Spacing - S 45 x D 37 x D 34 x D Type of Explosive

* Powder Factor: The powder factor is defined as the explosive necessary to fragment 1 m of rock. This equation can also be defined as the amount of explosives over the cubic yards of material desired to be blasted. Kg of explosive used/volume of material blasted. =kg/ m 3 . It is the opinion of many specialists this is not the best tool for designing a blast, unless it is referring to pattern explosives or expressed as energetic consumption. The size of the fragmented rock should match the bucket size of the excavator and also the grizzly size of the primary crusher. It can be also expressed in ton/kg. The following figure shows, how the total operating cost varies with the powder factor.


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A relation of average fragmentation size in function with burden and powder factor is shown in the figure below:

* Charge Distribution: The required energy needed to produce rock breakage is not uniform in bench blasting. The energy generated by the explosive must overcome the tensile strength of the rock (section C D D’C’) and the shear strength (section A’B’C’D’). To achieve this effect the explosive with the greater density and strength should be placed on the bottom of the blasthole, known as the bottom charge. It should be noted that placing this charge on the bottom of the blasthole increases the diameter of shaped charges by roughly 10%. The explosive with the lighter density should be placed in the column; this is known as the column charge.



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The energy per unit length for the bottom charge should be roughly 2 to 2.5 times more then the energy necessary for rock breakage. Recommended lengths of bottom charges are given in Table

Design Parameter Bottom charge length

Variation of bottom charge length with UCS & Diameter Compressive strength (MPa) Soft Medium Hard < 70 70-120 120-180 30 x D 35 x D 40 x D

Very Hard > 180 46 x D

The height of the column charge is calculated by the difference between total lengths of blast hole and the sum of stemming and bottom charge lengths. For further read: * Factors in designing of blasts, flyrock, industrial explosives used and safe operation of bulk explosives in opencast mines; Journal of Mines, Metals & Fuels; September 2010 (pp. 255 - 261) OR Clik: [Factors in Designing of Blasts, Flyrock, Industrial Explosives used and safe operation of Bulk explosives in opencast mines] * Blast Fragmentation Appraisal-Means to Improve Cost Effectiveness in Mines ----------------------------------------------------------------------------------------------------------Partha Das Sharma’s Bio-data:


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Partha Das Sharma (P.D.Sharma) is Graduate (B.Tech – Hons.) in Mining Engineering from IIT, Kharagpur, India (1979) He has very rich experience both in Mining operation and Marketing / Export / offering of Technical Services of Explosives, ANFO, Bulk explosives, Blast designing etc. Visited number of countries in Africa, South East Asia etc. He was associated with number of mining PSUs and explosives organizations, namely MOIL, BALCO, Century Cement, Anil Chemicals, VBC Industries, Mah. Explosives, Solar Explosives before being a Consultant. He has presented number of technical papers in many of the seminars and journals on varied topics like Overburden side casting by blasting, Blast induced Ground Vibration and its control, Tunnel blasting, Drilling & blasting in metal underground mines, Controlled blasting techniques, Development of Nonprimary explosive detonators (NPED), Hot hole blasting, Signature hole blast analysis with Electronic detonator, Acid Mine Drainage (AMD), Mining and Industrial dust etc. TECHNICAL PAPERS PRESENTED IN SEMINARS/JOURNALS: * Overburden Blast Casting with SMS Explosives – A case Study, Special Issue on Explosives & Blasting, Indian Mining & Engineering Journal, November 1998. * Blast Casting with SMS – A case study at Sasti Opencast mine, “Visfotak” ‘98, National Seminar on Explosives, Nagpur (India) * Control of adverse effects of Explosives Blasting in mines by using Shock tubes (Non-electric) Initiation system and its Future challenges; Advances in drilling and blasting techniques- Procc. of DRILL BLAST ’99 – National Seminar on drilling and blasting, Bhubaneswar, (India) January 2000. * Overburden side-casting by blasting – An effective way of reducing operating cost in large opencast mines; Journal of Mines Metals and Fuel, November 2004 (Sp., issue on development in surface mining technology – Calcutta, India). * Overburden side-casting by blasting – Operating Large Opencast Coal Mines in a cost effective way; Procc. of 1st Asian Mining Congress - Asian Mining: Towards a new resurgence (Vol. I), Seminar organised by MGMI at KOLKATA (India) from 16th – 18th January 2006 (pp. 307 – 315). * Non-Primary explosive detonator (NPED) – An eco-friendly initiating system for commercial blasting is the need-of-the-hour for Indian mines; Journal of Mines Metals and Fuel, March 2006. * Open pit blasting with in-hole delays and / or pre-splitting of production blast – Measures to control adverse impact of complex vibration arising due to presence of underground workings in the vicinity or in otherwise sensitive areas; Mining Engineers’ Journal, August 2006.


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EMPIRICAL EQUATIONS SUPPORTING BLAST DESIGNS Discussion on Design Parameters and Related Matters * Tunnel blasting – emulsion explosives and proper blast design are the pre- requisite for better efficiency; Journal of Mines Metals and Fuel, September 2005. * Improved Blasting technique is the key to achieve Techno-Economics of high production Underground Metalliferous mines; Indian Mining & Engineering Journal, December 2006. * Enhancement of drilling & blasting efficiency in O/C & U/G mines – Use of modern precision drilling, electronic delay detonator system and other sophisticated equipments with new generation emulsion explosives are the need-of-the-hour; Mining Engineers’ Journal, February 2007. * Improved Blasting with precision drilling patterns in Underground Metalliferous mines; Procc. ‘Golden Jubilee Seminar’ on Present status of Mining and future Prospects, organized by MEAI (6th to 8th April 2007) at Hyderabad, India. * Reduction of Ore dilution/Ore loss in underground metalliferous mines, lies on mitigation of blast induced vibration to a great extent; Mining Engineers’ Journal, August 2007. * Controlled Blasting Techniques – Means to mitigate adverse impact of blasting in Open pits, Quarry, Tunnel, UG metal mines and construction workings; Mining Engineers’ Journal, January 2008. * Controlled Blasting Techniques – Means to mitigate adverse impact of blasting; Asian mining: Solutions for development and expansion (Vol. II), Procc. of 2nd Asian Mining Congress, organized by MGMI at Kolkata (India) dt. 17th – 19th January 2008 (pp. 287 – 295). * ‘Electronic detonators – An efficient blast initiation system, Mining Engineers’ Journal, India, October 2008. * ‘Electronic detonators – Results in substantial techno-economic benefits for large mining operations’, Mining Engineers’ Journal, India, February 2009. * Innovative “Signature-Hole Blast Analysis” Technique to predict and control ground vibration in mines; Asian mining – Resurgence of mining in Asia: Prospect and challenges, Vol. II (pp. 211 – 223), Proceedings of 3rd Asian Mining Congress (22nd – 25th January 2010, at Kolkata, India), Organised by MGMI, Kolkata. * Charging and blasting in hot strata condition in opencast coal mines: identifying crucial aspects for effective safety management; Journal of Mines, Metals & Fuels; India; January – February 2010; (pp. 21). * Techniques of controlled blasting for mines, tunnels and construction workings – to mitigate various blast induced adverse effects; Journal of Mines, Metals & Fuels; June 2010 (pp. 152-161). * Factors in designing of blasts, flyrock, industrial explosives used and safe operation of bulk explosives in opencast mines; Journal of Mines, Metals & Fuels; September 2010 (pp. 255 - 261). * Acid Drainage in Mines, African Mining Brief Online Jan - Feb 2011, (http://www.ambriefonline.com/jan-feb11%20guest.html), Acid Mine Drainage (AMD) Author’s Published Books: 1. "Acid mine drainage (AMD) and It's control", Lambert Academic Publishing, Germany, (ISBN 978-38383-5522-1). 2. “Mining and Blasting Techniques”, LAP Lambert Academic Publishing, Germany, (ISBN 978-3-8383-7439-0). 3. “Mining Operations”, LAP Lambert Academic Publishing, Germany, (ISBN: 978-3-8383-8172-5).


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EMPIRICAL EQUATIONS SUPPORTING BLAST DESIGNS Discussion on Design Parameters and Related Matters 4. “Keeping World Environment Safer and Greener”, LAP Lambert Academic Publishing, Germany. ISBN: 978-3-8383-8149-7. 5. “Man And Environment”, LAP Lambert Academic Publishing, Germany. ISBN: 978-3-8383-8338-5. 6. “ENVIRONMENT AND POLLUTION”, LAP Lambert Academic Publishing, Germany. ISBN: 978-3-83838651-5 Currently, author has following useful blogs on Web: • http://miningandblasting.wordpress.com/ • http://saferenvironment.wordpress.com • http://www.environmentengineering.blogspot.com • www.coalandfuel.blogspot.com Author can be contacted at E-mail: [email protected], [email protected], ------------------------------------------------------------------------------------------------------------------Disclaimer: Views expressed in the article are solely of the author’s own and do not necessarily belong to any of the Company. ***


Empirical Equations Supporting Blast Designs Discussion on Design Parameters and Related Matters

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