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By Hesam Dehghani, Farid Alimohammadnia, Nima Babanouri, Mokhtar Kalhori

In open-pit mining, blasting pattern design is performed to achieve a desired degree of fragmentation for loading, hauling, and crushing operations. Inappropriate fragmentation of rock will result in numerous problems such as incurring extra costs for secondary blasting, production interruption, etc. Wetness of the blast holes is one of the most important factors, which can cause the inappropriate rock fragmentation. Lowering ambient temperature in the vicinity of the explosion, the presence of water in the hole adversely affects the quality of the explosive while dissipating part of the explosion energy, ending up with inappropriate fragmentation problems. In the present research, in order to study blast-induced fragmentation conditions in Baba-Ali Iron Ore Mine, three scenarios were considered: explosion of ammonium nitrate with fuel oil (ANFO) in dry hole, explosion of ANFO in wet hole, and explosion of the emulsion in wet hole. For this purpose LS-DYNA, a numerical simulation code was used. Results of the present research showed that, compared to the explosion of ANFO in wet hole, the explosion of the emulsion produced more uniform distribution of fragment size, while the explosion of ANFO in dry hole resulted in fewer large fragments.

By Hesam Dehghani, Farid Alimohammadnia, Nima Babanouri, Mokhtar Kalhori

In open-pit mining, blasting pattern design is performed to achieve a desired degree of fragmentation for loading, hauling, and crushing operations. Inappropriate fragmentation of rock will result in numerous problems such as incurring extra costs for secondary blasting, production interruption, etc. Wetness of the blast holes is one of the most important factors, which can cause the inappropriate rock fragmentation. Lowering ambient temperature in the vicinity of the explosion, the presence of water in the hole adversely affects the quality of the explosive while dissipating part of the explosion energy, ending up with inappropriate fragmentation problems. In the present research, in order to study blast-induced fragmentation conditions in Baba-Ali Iron Ore Mine, three scenarios were considered: explosion of ammonium nitrate with fuel oil (ANFO) in dry hole, explosion of ANFO in wet hole, and explosion of the emulsion in wet hole. For this purpose LS-DYNA, a numerical simulation code was used. Results of the present research showed that, compared to the explosion of ANFO in wet hole, the explosion of the emulsion produced more uniform distribution of fragment size, while the explosion of ANFO in dry hole resulted in fewer large fragments.

By Ananta L. Yennamani

Blast movement measurement (BMM) transmitters are a practical method to measure three-dimensional (3D) rock movement because of blasting. The BMM transmitters are activated, programmed and installed in specially designed drill holes prior to the blast and the BMM detector locates the transmitters after the blast. The BMM software then calculates and summarizes the 3D movement of each BMM transmitter ball. This information helps to redefine the ore and waste boundaries and enables improved ore and waste selection, resulting in a genuine step change in grade control. The BMM program was carried out at Newmont Mining Corp.?s Phoenix Mine, in northern Nevada near the town of Battle Mountain. The mine is in a historic mining district that contains gold, silver and copper. The geology of the pit consists of different formations made up of sandstone, siltstone, limestone, chert-pebble conglomerate, shale, skarn and quartzite, to name a few. Some of the rock is extremely hard and a powder factor of up to 0.5 kg/t (1 lb/st) is required to get adequate fragmentation to ensure sufficient throughput through the crusher and the mill. The result of using such a high powder factor translates into considerable blast movement and heave. This movement means that the ore/waste digging polygons estimated by the ore control engineer from sampling the initial blast holes will no longer be located in their original posi-tions. The purpose of this project was to test the measurement method and to see if locating the digging polygons in the broken muck pile and excavating to those polygons will improve grade control through the mill. Case studies (Zhang, 1994), (Taylor, 1995), (Gilbride, 1995), (Harris, 1997), (Taylor, 2003), Barrick Goldstrike (Gold-strike Mine, 2008), Ruby Hill Mine (Hilkewich, 2009), (Porcupine Mine, 2005) and (Aguirre, 2010) have shown that, by accounting for the blast movement, there is a potential to increase mineral recovery by as much as 25 percent for individual blasts for only a modest cost increase.

Jan 1, 2011

By Marcel Laflamme, Sebastián Ibarra‑Gutiérrez

Improvement on controlling the stope limits has enabled the assessment of technical performance in several underground mining operations. Technical performance, integrated to stope geotechnical and geomorphological parameters, constitutes the basis which guides mining engineers in the Mineral Reserves Estimation process. Moreover, there are established definitions to report this information in Canada, incorporated in NI 43–101, which stipulates that “A Mineral Reserve includes diluting materials and allowances for losses that may occur when the material is mined” [1]. However, some loopholes have been identified regarding practical standards followed to specify the way to include and report ore losses when the lost material was blasted and must be left behind unrecovered inside the stope. In fact, general efforts to estimate and quantify the negative effects of operational ore losses only consider under-break ore loss. However, mining engineers frequently confront situations requesting the anticipated closure of a stope. Therefore, reconciliation process should reflect the amount of unrecovered Mineral Reserve. This work presents a discussion of possible practices to integrate adequately operational ore losses consisting of blasted but not recovered material and focuses on how different approaches regarding the material constitution could severely affect the stope technical performance interpretation.

Aug 9, 2021

By D. Donison

The practice of blasthole mining is typically applied in competent ground conditions, primarily because of the open stope dimensions. At Magma Copper -Superior Mine, a field pilot project was completed examining a variation of blasthole mining in poor ground conditions. The results were very encouraging not only from a technical viewpoint, but also productivity viewpoints. This paper describes the pilot project and the resulting implications.

Jan 1, 1994

By H. Peter Knudsen

In most open pit gold mines, ore-waste selection is based on estimates made from blasthole samples. These estimates frequently are biased. This paper focuses on the source of the conditional bias. While it is true that high grades and "outliers" can induce a bias, they may not be the only reason for the persistent bias seen at many mines. In this paper, it is shown that the poor blasthole sampling can cause a conditional bias in the estimates.

Jan 1, 1993

By H. P. Knudsen

In most open pit gold mines ore waste selection is based on estimates made from blasthole samples. These estimates frequently are biased. While it is true that high grades and "outliers" can induce a bias they may not be the only reason for the persistent bias seen at many mines. This paper focuses on a source of the conditional bias. In this paper it is shown that the poor blasthole sampling can cause a conditional bias in the estimates.

Jan 1, 1991

By Louis P. Gignac

The Lac Shortt gold deposit is located in Northwestern Quebec, at some 115 km west of the Opemiska Division and 380 km northeast of the Lac Dufault Division, as shown on Figure 1. Falconbridge's exploration crews drilled the first six holes in the gold-bearing ore zone in 1979. After further diamond drilling from surface in 1980, the decision was taken to carry out an underground exploration program. A three-compartment shaft was sunk to a depth of 340 metres in 1981 and 1982; it was followed by an extensive exploration program, which included 1073 metres of cross-cutting and footwall drifting, 645 metres of ore drifting and 11 200 metres of diamond drilling. The underground work was completed by mid-winter 1983. It was followed by a complete feasibility study of the Lac Shortt Project. After some $Can 14.2 millions in exploration expenditures, the Board of Directors of Corp. Falconbridge Copper approved in July 1983 additional expenditures of $Can 45 millions to bring the Lac Shortt Mine in production at the capacity of 261 000 tonnes per year. Gold price was $US 425 per ounce at the time.

Jan 1, 1985

By Peter N. Blakey

The Kidd Creek orebody is a massive, base metal deposit with widths up to 600 feet and a strike length of 2,200 feet. The orebody dips eastwards at 70° to 80° and is amenable to sub-level open stoping. Five distinct zones (Fig.1) based on structural wall rocks and ore type parameters occur within the orebody. These zones are: Zone 1 - Mainly copper stringer ore Zone 2 - Massive pyrite ore Zone 3 - Massive sphalerite and chalcopyrite Zones 4 and 5 - Are banded and variable in nature Compressive strengths vary from zone to zone with an average of 22,000 p.s.i. and highs of 40,000 p.s.i. The mine is developed from a ramp extending from surface to the 1600 foot level and a 24 foot diameter, 3050 foot vertical shaft with main levels at 400 foot intervals. Sub-levels at 100 foot vertical intervals are driven from the ramp. transverse and longitudinal pillars to support the crown pillar left under the open pit. The layout allows an opportunity to enlarge or reduce pillar sizes and also to introduce fill on any horizon should conditions warrant. It also offers a large number of operating stopes spread through the various ore zones giving good ore blending characteristics. All ore from the stopes is handled through drawpoints using Wagner ST8 scooptrams with an average one way haul length of 650 feet.

Jan 1, 1976

By Joseph S. Malesky

The discovery and development of explosives mark one of the most important findings in the history of civilization. Without explosives our vast economic enterprise concerning the mining of coal, copper, iron, aluminum, and various other ores could not function. Explosives are used in all kinds of construction works, in the razing of old buildings, blasting of trenches, driving of tunnels, and in many other instances common to our everyday life. For these reasons explosives are rightfully referred to as the forerunner of progress. An explosive may be defined as a substance (usually a mixture) that, when flame or shock is applied, yields readily to rapid combustion or oxidation, accompanied by the formation of such relatively large quantities of gases as to produce much violence and pressure which, in turn, react against their surroundings and cause what is termed an explosion. By their very nature the action of explosives is destructive and disruptive of adjacent materials. Decomposition of explosives can occur through deflagration, which is rapid burning or combustion, as in the case of black powder, or through detonation, which is an almost instantaneous disruption, as in the case of dynamites and blasting agents.

Jan 1, 1981

By G. T. Austin

Every year the United States consumes millions of dollars worth of abrasive materials as blasting media. Entrained in either a gas or liquid stream or propelled by paddles or wheels, they are directed onto the surface of another material to modify that surface. The modifications to impacted surfaces include cleaning, cutting, deburring, etching, finishing, and peening. Blasting media can be either organic or mineral abrasive materials. Information and data presented are for only a portion of the mineral abrasives, they are; fused aluminum oxide, coal slag, copper slag, garnet, metallic shot and grit, nickel slag, olivine, silica sand, silicon carbide, and staurolite.

Jan 1, 1994

By C. L. Barron, M. L. Smith, K. A. Prisbrey

SAG mill throughput is often a bottleneck to increasing plant productivity. Improved size distribution control in the mill feed can be used to increase mill production. This can be accomplished by controlling blast fragmentation as a function of energy input and rock characteristics. Such an approach requires the integration of rock fragmentation and comminution modeling. The optimal mill feed size distribution is determined by adapting a population balance comminution model to a measure of rock hardness or fragility, which is also related to the rock fragmentation that is achieved by blasting. The relationship between the size distribution and fragility of the feed and the mill's product and the critical size in a SAG mill is examined.

Jan 1, 1994

By M. L. Smith

SAG mill throughput is often a bottleneck in increasing metal productivity. Improved size distribution control in the mill feed can be used to increase mill production. This can be accomplished by controlling blast fragmentation as a function of energy input and rock characteristics. Such an approach requires the integration of rock fragmentation and comminution modeling. The optimal mill feed size distribution is determined by adapting a population balance comminution model to a measure of rock hardness or fragility which is also relevant to the rock fragmentation that is achieved by blasting. The relation between the size distribution and fragility of the feed and of the mill's product and the critical size in a SAG mill is examined.

Jan 1, 1993

By Lewis L. Oriard

INTRODUCTION In recent years, there has been a trend in the direction of larger drilling equipment and larger diameter blastholes. Although this change has improved the efficiencies and reduced the costs in many operations, it has increased the potential for damage to underground openings. In addition, in many instances one now finds more sophisticated delicate instruments, automated control facilities, and a large variety of structures in proximity to blasting activity. The combined effect of larger-scale blasting activity and its proximity to various features of interest is such that there is an increased need for a more refined analysis of blasting effects and their control. BLASTING EFFECTS ON ROCK SURFACES The Breakage Mechanism In order to develop techniques for controlled blasting, one must first understand the features of the mechanisms by which blasting causes rock breakage to occur. These features have not been easy to demonstrate, mostly due to the difficulty in making tests and observations at the high stress levels and short time durations involved. When an explosive charge is detonated, the material surrounding the charge is subjected to a nearly instantaneous, very high pressure [on the order of 1.4 to 13.8 GPa (0.2 to 2.0 X 106 psi), depending on the explosive]. If the charge is coupled to "average" rock, this pressure will pulverize the surrounding rock for a distance on the order of 1 to 3 charge radii in hard rock, and to a greater distance in softer rock (this is also dependent on the type of explosive). As the pressure wave passes into the rock, high tangential stresses cause radial cracks to appear, and the nearly discontinuous radial stress zones gen¬erated by the shock front may cause tangential cracks to appear. The extent of these cracks depends on the energy available in the explosive, how quickly the energy is transmitted to the rock, and the strength properties of the rock. The discontinuous shock front is quickly dis¬sipated, but the expanding gases generate a longer-acting pressure. A compressive pulse travels to the nearest face or internal rock boundary where it is reflected in tension. The tensile strengths of most rocks are roughly 40 to %o of their compressive strengths, so the rock may now fail in tension whereas it may have been able to support the diminished compressive phase without failure. The ten¬sile deflection typically produces a failure described as tensile slabbing or scabbing. Laboratory experiments and field experience have pretty well established that several mechanisms are involved. These include (1) the classical case of tensile parallel slabbing when the pressure pulse is reflected at a free surface; (2) failure under quasi-static compressive loading (the shape is normally irregular due to discontinuities in the rock); (3) radial cracking under the action of tangential stresses at the periphery of the expanding pressure pulse; (4) peripheral cracking at the discontinuous shock front which is quickly dissipated; and (5) additional mass shifting due to the venting of the explosive gases. The first three items have received much attention in the laboratory and the literature. The complex effects of gas venting are difficult to test in the laboratory because of the difficulty in reproducing the many weak planes and discontinuities typical of most field conditions, which play such a prominent role in determining the behavior of the rock mass subjected to blasting. Unfortunately, gas venting effects can be pro¬jected to significant distances under certain field conditions, and are sometimes difficult to control. It is not unusual for gas venting to be the overriding factor in determining the final geometric shape and physical condition of the finished excavation. Sources of Damage For the purposes of this discussion, damage includes not only the breaking and rupturing of rock beyond the desired limits of excavation but also an unwanted loosening, dislocation, and disturbance of the rock mass the integrity of which one wishes to preserve (such as mine pillars, underground openings, etc.). The sources of damage include, of course, all those physical features of the rock breakage mechanism. Each of these effects must be limited to the desired zone of breakage and excavation if the integrity of the remaining rock mass is to remain undiminished. The primary zone of rock breakage usually can be controlled in the normal process of field experimentation to determine proper charge sizes and location for primary excavation. However, it frequently happens that there is damage from sources which are more difficult to account for in the design process, which are often overlooked. These are (1) the overbreak due to poor drilling control, (2) dislocation of rock (mass shifting) due to venting of explosive gases, and (3) loosening or dislocation due to the influence of seismic waves (ground vibrations). CONTROL OF ROCK BREAKAGE Importance of Geometry In studying the rock mass and blasting design con¬siderations, it is important to keep in mind the geometric relationships among charge size, shape, and position, and the physical features of the rock mass to be preserved. The features of principal interest are the external shape and position of the rock mass relative to blasting, and the position and attitude of any weak planes in the rock mass. The Sequence of Blasting and Excavation Events Unfortunately, there are too many times when the task of preserving delicate rock is considered hopeless, and because of this attitude, no further effort is ex¬pended towards caution or control. In such cases there is often a failure to recognize the importance of the se¬quence of the procedures. Attention to this can greatly reduce unwanted effects at minimum cost. Perimeter Control The requirements for perimeter control are highly dependent on the special needs of each particular proj¬ect. The desirable degree of control is a highly variable

Jan 1, 1982

By E. De Souza, H. Heidrich

In underground operations the evaluation of blasting fume dilution times and the application of recirculation techniques require accurate determination of the concentration of toxic fumes produced by the explosive used. This paper describes a methodology for obtaining concentrations of detonation products from commercial explosives tested under typical rock confinement and looks into the development of techniques for reducing toxic gases produced by explosives. This paper also analyses the application of techniques for the removal of toxic gas which would enable reduced re-entry times, a reduction in the quantity of fresh air and the use of air recirculation techniques, thereby providing substantial economic benefits to the industry. INTRODUCTION Toxic fumes from blasting operations are among the most common and serious hazards encountered underground. The implication of high concentrations of toxic gases emitted from blasting on the mine ventilation system are very significant and, in many cases, catastrophic results have occurred. Because blasting is a routine practice underground, the mining engineer is always confronted with the problem of controlling the fumes within acceptable levels. Typical control procedures employed underground include preventing or reducing the amount of fumes liberated from blasting through the proper selection of explosives and proper blasting techniques; removing the fumes through local exhaust systems or auxiliary ventilation (dilution ventilation); blasting off-shift, or at restricted times; localizing or isolating the effects of blasting where workings are isolated or sectionalized in the ventilation system; and absorbing some of the components of the fumes using air-water sprays. Dilution is one of the most effective control measures against blasting fumes and is universally applied. However, for precise determination of the dilution times for the fumes to reach safe concentrations, accurate determination of the concentration of toxic fumes produced by the explosives used is required. Research work has been carried out in determining the types and volumes of gases produced from blasting. In the case of ideal detonations, the nature and quantity of the fumes produced can be predicted by thermo-hydrodynamic calculations. However, commercial explosives exhibit non-ideal behaviour which is enhanced by the relatively small diameters of application in underground mines. The confinement offered by the surrounding rock further complicates the situation. It is well recognized that confinement affects performance and consequently fume generation. A detailed data base of fumes generated by commercial explosives at various conditions is introduced and the effect of blasting parameters such as, confinement, shock desensitization and incomplete detonations on toxic fume production is described for a variety of commercial explosives in use today. Thermodynamic predictions on fume production by the explosive compositions tested are also presented and compared with experimental results and their significance in assessing the quantity of fumes produced is discussed. A mathematical model capable of determining dilution ventilation requirements and dilution times to reduce the fume concentrations in the blasting zone to acceptable levels has been developed and it is being incorporated into a network system so that the effects of blasting can be simulated in typical mine ventilation systems. Solutions to the problem of controlling blast fumes within acceptable levels are described. These solutions include the use of control procedures which prevent and reduce the amount of fumes liberated from blasting. The use of low fume explosives, proper blasting techniques, dilution ventilation and filtration can be combined to effectively control fume production and save on costs. The application of these techniques to enable recirculation of underground air with reduced quantities of fresh air requirements is evaluated in a case study.

Jan 1, 1993

By L. Donovan Clark

The term "powder factor" is almost universally accepted as the measure of the efficiency of a blasting program. It represents the economics of blasting since it designates quantity of explosive (thereby providing quality control in rock fragmentation) in terms of the associated volume broken. It is the wide use of this concept that explains the interest shown in an approach to rock blasting preferred by A. W. and Z. W. Daw* in a treatise on 3 blasting (1) substantiating its use in the form, Q = CVW3 for single shot hole, where Q = quantity of explosive lb Cv = blasting coefficient W3 = unit volume of rock It is obvious that the term Cv could also be called the apparent powder factor should the equation be written in the form, Cv = Q/W3 (lb per cubic unit). However, while the approach was ingenious** and the practical applications straightforward, their original, rather limited experimental work, along with certain assumptions, was questioned.

Jan 1, 1970

By Edward J. Karkoska

The film you see is a slow motion view of a typical half million ton taconite blast at one of Erie's mines. This blast is made up of 220 drill holes and a total of 325 thousand pounds of blasting agent. The second view of the same shot is taken by a remote control hi speed camera located 1500' directly in front of the blast. Shortly you will see some violent camera movement as the air blast shock wave strikes the camera location. As the blasting sequence progresses the rock fly becomes evident, and if you will look closely amidst the clouds of smoke, rock fragments weighing up to 1/2 ton begin to cascade into view. These missles are the chief concern of blast security. The next camera shot at 4000 frames/second vividly describes the primacord firing lines detonating the blast. Travelling at 22,000 ft./sec., this white snake of fire travels to each hole, detonating the blasting agent. Before we get into the matter of blast security, let me locate our operation for those of you not acquainted with Minnesota's iron ranges and tell you something about our company.

Jan 1, 1969

By C. J. Gilbert

The mining method at the Cannon Mine uses heading and bench stopes, which are subsequently backfilled with cemented high-strength fill prior to pillar removal. The dimensions of the pillars, 24 feet wide by G5 to 95 feet high, re-quire careful blasting in order to not damage the rock, compromising pillar strength and complicating later removal. Additionally, the proximity of dwellings makes vibration control a must for good-neighbor relations. A blast monitoring study was conducted in order to optimize bench blasting practices to maximize fragmentation, reduce costs, minimize well damage, and minimize ground vibration.

Jan 1, 1989

By Clark J. Gilbert

The mining method at the Cannon mine uses heading and bench stopes, which are subsequently backfilled with cemented high-strength fill before pillar removal. The dimensions of the pillars are 7.3 m wide by 19.8 to 28.9 m high (24 ft by 65 to 95 ft). They require careful blasting in order to not damage the rock, compromising pillar strength and complicating later removal. Additionally, the proximity of dwellings makes vibration control a must for good neighbor relations. A blast monitoring study was conducted in order to optimize bench blasting practices to maximize fragmentation, reduce costs, and to minimize wall damage and ground vibration. The Cannon mine is a joint venture between Asamera Minerals (US) Inc., a 92% owned subsidiary of Gulf Canada Ltd., and Breakwater Ltd. Asamera is the operator. The mine is located at the south end of Wenatchee, WA, a town of roughly 45,000 people on the eastern flanks of the Cascade Mountains. Gold is the primary product at the mine. The rock is extracted using a sublevel end-slicing method on 15 m (50 ft) lifts, with subsequent backfilling and pillar removal. The stope widths, dictated by rock conditions at 7.3 m (24 ft), are the same in the primaries as in the pillars, or secondaries.

Jan 1, 1990

By J. Y. Perez, H. M. Ewoldsen, L. L. Oriard

Near the village of Villarino de los Aires, in the province of Salamanca, Spain, is an existing underground hydroelectric power plant. A major enlargement was undertaken to increase the electrical generating capacity, under a contract awarded to a joint venture of Dragados & Construcciones, S.A., Entrecanales y Tavora, S.A., and S.A. Conrad Zschokke. The enlargement required the excavation of an underground system of tunnels, shafts and chambers adjacent to existing facilities and interconnected with these facilities. The existing machine hall and transformer chamber were both extended, requiring excavation of the rock that constituted the existing end walls of these chambers. The drilling, blasting and excavating of the underground system had to be done without damaging existing underground openings or any of the existing structures, equipment or instrumentation facilities. This required careful control of vibrations, air waves and dust. The joint-venture contractors preferred to use electric blasting methods, if these methods could be implemented in a safe manner. The safety question had to be evaluated carefully. High electrical potentials existed at the site because of the electrical generating and transmitting facilities that could not be shut down. Electrical fields were carefully studied prior to the beginning of the work. This was done both in the underground environment and above the ground surface. Special procedures were followed to ensure safety with

Jan 1, 1974