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By Brian L. Beninger

This paper discusses the application of the computer to coal preparation process predictions. In broad terms, the computer is used to perform two main functions. Firstly, the given float-sink analysis is interpolated and extrapolated to cover the entire actual range of raw coal specific gravity. Assumptions based on historical data must be made at times. Secondly, the partition curve, which, from historical data, defines the efficiency of a specific coal washing process, is applied to the washability data in order to determine the expected clean coal yield and quality. The vast amount of numerical work which is required to make a prediction of this type exhibits the need for computer methods in this area.

Jan 1, 1979

By W. J. Wright

Summary Statement The following notes on the geology of the Grand Lake coalfield are intended to be introductory to a second paper being read at this meeting, which paper will describe the earlier and the more modern mining methods. The coal industry of the Grand Lake coalfield is centred chiefly within a three-mile radius of Minto, but minor operations are dotted over an area of approximately 50 square miles extending easterly to the vicinity of Chipman and Coal Creek (see Figure 1). The coal comes from one seam of Pennsylvanian age. The seam varies in thickness from 16 to 30 inches and lies almost horizontal at depths of 0 to 125 feet. It has been mined and prospected over an area of approximately 150 square miles, but in a portion of this area the commercial possibilities have been destroyed by erosion. The industry is the oldest of its kind in North America, but it did not at tract particular publicity until about 1914. Since that date, mining methods have improved and the output has rapidly increased until, in 1934, it amounted to 271,335 long tons, and furnished employment for about 1,000 men (1). Approximately 73 percent of this amount was produced by the Minto Coal Company, the Miramichi Lumber Company, the Avon Coal Company, and Welton and Henderson, and 25 per cent by the Rothwell Coal Company, Harvey Welton Limited, G. H. King, the Reade Construction Company, and the Newcastle Coal Company. The remainder was produced by J. W. Carter, G. D. Crossman, T. H. Swift, C. S. Yeamans, and several other small operators. The coal is won by underground mining, except in the case of the Reade Construction Company, and a minor operation by Ashfield and Pieton, who have employed open stripping. It is mined only as required to fill definite orders, and is loaded directly into cars for shipment. In 1934, the larger mines did not operate full time, and the production could easily have been increased if markets had been available. The coal is sold in four grades: screened or lump, run of mine, slack, and stripping. The chief market is with the railways and commercial plants, with a fair amount for domestic purposes.

Jan 1, 1935

By D. A. Sloan

"THE FIRST COMMENTATOR that I will call on is Mr. W. G. (Bill) Brissenden, Vice-President, Mines, Noranda Mines Ltd. Bill, would you care to comment.W. G. BRISSENDEN ...Mr. Chairman, Ladies and Gentlemen, I spent some time reading the four papers that were presented here, yesterday evening, and decided that the four to six minutes that I had been allotted for this afternoon were going to be pretty tough to meet. It's a very big subject and I just hope that I will be able to stay within my time limit. As has been so aptly demonstrated by the members of our panel here this afternoon young men in growing numbers continue to be attracted to careers in the mineral industry, especially to careers in mining engineering. Now from 80 the point of view that the profession may have regained its status this is indeed welcome news. However, perhaps all of the members seem to have the suspicion that the number of mining engineers that the industry can absorb is a much lower figure than previously forecast. It is said that technical school graduates are now occupying some of the jobs once considered the private preserve of the mining engineer and also that mining engineering undergraduates are having trouble locating summer employment. If these interpretations of mine are correct then we are now confronted not only with our old dilemma - whither the mining engineer - but also with an urgent need to solve it. We cannot allow our profession to become once again an undesirable one in the eyes of our youth because if we permit such a thing to happen a second time the profession may never recover."

Jan 1, 1972

By P. S. Arthur, G. R. F. Alvear

Development of a new smelting process is risky. Many novel processes in the metals industry failed to fulfil the initial expectations of their inventors. Large sums of capital have been invested in technologies that either didn't provide a satisfactory return on investment, or didn't perform at all, and had to be replaced. One of the keys to success in process development is incremental scale up of the technology, with potential problems being identified and rectified as early as possible. In this way the cost of any rectification that is required can be minimized. The ISASMELTTM process was developed to maturity by Mount Isa Mines through the 1980's and 1990's using incremental scale up. Commercialization only occurred once the process had been proven on laboratory, pilot and demonstration scale over many years. As a result ISASMELTTM technology now operates successfully at numerous plants around the world. The methodical approach to development of the technology has allowed owners to modernise their existing operations or create new businesses with significantly reduced technical risk.

Jan 1, 2011

By S. Banisi, J. Marois, A. R. Laplante

"This paper compares the behaviour of gold and ore in the grinding circuit of Hemlo Mines Ltd. Particular attention is given to grinding kinetics, classification behaviour, and liberation. For grinding kinetics, laboratory grinding of gold flakes and silica, 840 µm to 1200 µm in size (14/20 mesh), was also performed to provide a standard for comparison and to study in a controlled environment the mechanisms of breakage, folding , smearing and impacting.Grinding kinetics were characterized at laboratory scale with breakage and selection functions. The breakage function showed that 75% of gold disappearing from the 840 µm to 1200 µm class reported to the 600 µm to 840 µm; this figure is much higher than for silica (46%) or most other minerals, and implies that comparatively fewer fine fragments are produced from the breakage of gold flakes, on account of gold's malleability. The selection function of silica was four times that of gold. At plant scale, the ratio of the selection function of ore and gold increased from 6 for at 50 urn to JOO pm to 20 at 500 um to 1000µm.Observation of original flakes and their fragments with the scanning electron microscope showed that the main mechanism of ""grinding"" from the 840 µm to 1200 µm to the 600 µm to 840 µm class was folding rather than actual breakage. This was confirmed by a measure of individual flake weight. The postulated breakage cycle is that relatively globular flakes become increasingly serrated through grinding; they can then regain their original shape through folding and reflattening into cylindrical and spherical shapes, as the majority does; however, some fragments will at this point undergo actual breakage. Photograhic evidence is presented to support the proposed mechanisms.Smearing of gold unto silica was observed when silica was ground after a gold, and embedding of silica into gold flakes when gold was ground after a silica. Both mechanisms occur whenever gold and harder minerals are jointly ground, and result in ""unliberation"" of gold. It is postulated that smearing of gold onto other minerals plays an important role in explaining how gravity recovery can increase over-all gold metallurgy, especially when used ahead of flotation.Classification of gold in the primary hydro cyclones takes place at a much finer cut-size than ore, 20 µm vs 57 µm. This and the slower grinding kinetics result in high gold circulating loads in the primary loop, 6700%, and a gold content in the primary cyclone underflow (PCUF) is twenty-four times that of the ore. The gold size distribution for primary cyclone overflow is much finer than that of the ore, although they are essentially similar in the PCUF.Liberation, or more accurately ""gravity recoverability"", was assessed with a 7.5 cm (3 in.) Knelson concentrator. It varied between the thirties for the primary mill discharge and two cyclone overflows, to highs of 89% and 92% for the PCUF and secondary mill discharge."

Jan 1, 1991

By Roméo Roussin

"The yearly production from the Asbestos Hill pit is 1,550,000 tons of asbestos ore, with a yield varying between 18 and 20 % . The orebody was exposed at the top of a hill, and initial production started without having to remove any overburden. The projected ratio of barren rock to ore, based on the present orebody, was 2 to 1. However, as it was necessary to push back the upper benches, this ratio has changed to 3 to 1. The final pit will be 2,300 feet long by 1,000 feet at the longest and widest points. The final depth will be 680 ft. A pit slope stability study was carried out and an average wall angle of 60-70 degrees must be maintained. Preshearing of final pit limits will produce good wall stability. Permanent access ramps will be 60 ft wide, with an 8 % maximum grade. Geological conditions and the low rock temperature are the major causes of the operation's drilling and blasting problems."

Jan 1, 1976

By D. M. MacLean

THE property of Howey Gold Mines, Limited, is situated at Red Lake, Ontario, one hundred and seventy-five miles east of Winnipeg and eighty miles north of the Canadian National railway. The mine has been a continuous producer since April, 1930, and, at present writing, a total of 3,042,690 tons has been hoisted, giving a production of $6,422,858.55, for a grade of $2.32 per ton. From 1,064,477 tons mined during the past two years, however, the yield has been only $1,389,607.03, or $1.41 per ton (gold at $20.67). Present output is at the rate of about 1,470 tons. per day. THE ORE-BODY AND WALL-ROCK: The Howey ore-body is part of a quartz porphyry dyke, which has. intruded an area of greenstone. The dyke strikes north 65° east, and its dip varies from 85° to the south at the surface to practically vertical below the. 1,000-foot horizon. Both along the strike and on the dip, the width of the dyke varies from 25 to 200 feet, the widest portion occurring below the 1,000-foot level. To the east of the ore-body, the dyke increases in width and finally splits into tongues of porphyry enclosed in greenstone, diorite, and congloJ11erate. To the west, it passes into the adjoining property, its width at the boundary being 100 feet. The rock in the ore-body is hard to drill, but, because of numerous horizontal slip-planes, it breaks well when blasted. This characteristic makes it possible to break large breasts or benches with a relatively small number of holes. The ore broken in this manner is consequently quite coarse, a large part of it requiring secondary breaking.

Jan 1, 1938

By J. I. Carroll, J. A. Walker

"Abstract - Camelback is a small (=200 000 tonnes), moderate grade (5%-7% Zn+Pb), volcanogenic massive sulfide deposit, which occurs within the Nepisiguit Falls Formation of the Ordovician Tetagouche Group. The host rocks are tuffaceous sedimentary rocks (Little Falls member) that overlie quartz-feldspar porphyritic tufflavas (Grand Falls member). The hanging-wall sequence comprises rhyolite of the Flat Landing Brook Formation, overlain by the Forty Mile Brook tholeiitic basalt. A dike of unaltered andesite was intersected beneath the massive sulfides, but was not found in the hanging-wall sequence in the vicinity of the deposit. The stratiform part of the deposit is made up of two, steeply south dipping, subparallel massive lenses that average approximately 4 m in thickness. Each lens is zoned, with an upper part consisting mainly of a pyrite and a basal part containing pyrite, sphalerite, and galena, with Ag values. The Au content in the massive sulfides is low (average = 52 ppb for 271 samples), but tends to be enriched in the massive pyrite near the top of each lens. Oxide facies iron formation is spatially associated with the massive sulfides, and has been traced up to 1800 m along strike. Footwall hydrothermal alteration is typical of VMS systems, and is characterized by depletion in K2O and Na2O, and enrichment in MgO and Fe2O3T with proximity to massive sulfides. The massive lenses are underlain by intensely chloritic, fine-grained, tuffaceous sedimentary rocks containing locally significant sulfide (chalcopyrite > pyrite > pyrrhotite) veins, which are interpreted to represent feeder zone mineralization. Unlike many VMS deposits, there is no evidence of silicification beneath the massive sulfides at the top of the feeder zone. The oxygen isotopic compositions (d18O = -1.5‰ to -0.3‰) of hydrothermal chlorite coupled with sulfur isotopic compositions (d34S ˜ 12‰) of the massive sulfides suggest that seawater was the dominant fluid in the hydrothermal system. Normalized rare earth element diagrams show flat Ce profiles and positive Eu anomalies, indicating that modified seawater was involved in ore formation; i.e., high-temperature (=350ºC), acidic, and reduced prior to entrainment into the hydrothermal cell. Although, seawater was the dominant fluid in the sulfide-forming system, d34S values of stringer zone sulfides (8.5‰) coupled with elevated Sn (400 ppm) in one massive sulfide sample indicate that there was at least some magmatic component in the hydrothermal cell."

Jan 1, 2006

By D. Jean Hutchinson, Mark Diederichs

"Cable bolts are an important support component in modern underground mines. If an optimized cable bolt system is to be achieved, it is insufficient to focus on cable bolt layout design or on cable bolt installation in isolation. Cable bolt support, like all engineered systems, requires a holistic approach, encompassing integrated elements of design, implementation and verification. The “Cable bolting Cycle” described in this paper incorporates these components along with related tools and procedures which should be used together to ensure effective and efficient cable bolt support.Design involves the consideration of demand, capacity and constraints as outlined in this paper. The demands imposed on the sup-port system by the rock mass response to excavation under the combined action of stress and gravity must be matched to the capacity of the cable bolt array. The cable bolt system capacity is a function of the component strengths of the steel strand and the bond (load transfer) and by the deformation response and retention characteristics of the system. The design is bounded by operational and economic constraints.Implementation of the cable bolt array can then proceed and must include appropriate protocols of communication, training and quality control, in addition to both general and site specific installation procedures. Implementation will be discussed as the sum of these components.Implementation of the cable bolt array can then proceed and must include appropriate protocols of communication, training and quality control, in addition to both general and site specific installation procedures. Implementation will be discussed as the sum of these components.The information presented in this paper represents a very abbreviated version of much broader concepts, procedures and design charts, presented in a practical cable bolting handbook by Hutchinson and Diederichs (1996)."

Jan 1, 1996

By T. C. Holmes

"Dome mine, in the Porcupine mining camp of northern Ontario, is one of the important gold mines of Canada, its production of gold from the commencement in 1910 to the end of 1946 having a value of $149,000,000.GENERAL GEOLOGYThe oldest rocks in the mine are greenstones of Keewatin age(l). Those to the south are known as the southern greenstones, but little is known of their structure. Those in the northwest part of Fig. 1 and the north part of Fig. 2 are known as the northern greenstones or greenstone 'nose'. They face north, dip steeply north, and end to the east against an unconformity which pitches northeast about 20°. The volcanics are overlain unconformably by Timiskaming .conglomerate and slate. All these rocks are cut by porphyry. Closely associated with the porphyry is a group of rocks that have suffered extreme alteration. Diabase dykes intrude all the other rocks and the ore. The principal massive rocks in the mine are two adjacent flows known asthe andesite (south) and dacite (Fig. 1). They are composed chiefly ofmassive, medium-grained lava. Pillows, flow lines, and other typically volcanicfeatures are found along the tops of both flows and at other places in theandesite. These parts nearly everywhere show a slight schistosity panllelwith the flows except in the orebodies, where they are massive."

Jan 1, 1954

By F. E. Millett

The following notes deal chiefly with changes from steam to electric drives, or, in the case of the Bellevue pumps, from air to electric. Where electric power can be purchased or generated cheaply it can be used to advantage and is generally more economical than steam, especially in places remote from the central power-plant. Nothing could be more simple than the operation of electrical gear. It does not require a competent engineer constantly on the alert, oiling and watching bearings, and maintenance is at a minimum. The ease with which the current is transmitted from its source to where it must do its work is one of its greatest advantages. Transmission losses are very low. Generally speaking, the smaller the plant the lower the loss and 90 per cent, and better, efficiency is easily attained. In and around Bellevue 2300-volt, three-phase, is the standard distributing pressure; our power requirements and the distances transmitted, for the present, do not require more than this. Most of the heavier motors are 2300-volt .machines, which eliminates the cost and possible troubles of step-down transformers, and makes for a cheaper and less bulky motor. This applies to two fan, two hoist, and two compressor motors, the smallest being 75 and the largest 520 horse power. The standard pressure on power lines to all inside pumps is 550 volts, and 220 volts is general for the the tipple. While 110 volts is used for house and office lighting service, it is not economical around and in the mine. No matter what kind of device is used to lock and protect the lamps, they are invariably missing.

Jan 1, 1926

By K. Mirkhani, P. Gebski, W. Ying

The campaign life of an industrial metallurgical vessels strongly relies on the quality and quantity of the refractory lining. An operating furnace or converter refractory lining can be assessed either by thermal modeling or by acousto ultrasonic-echo (AU-E) technique or a combination. AU-E is a Non-Destructive Testing (NDT) stress wave propagation technique that has been utilized on a wide variety of smelting vessels for over 12 years with good success. AU-E can determine the thickness of remaining refractory lining, estimate the thickness of build-up or accretion, detect and locate metal penetration and impregnation into the lining, and identify hydrated or cracked areas of refractory. This paper details the AU-E principles, equipment and recent improvements to the data analysis to further improve speed and accuracy. In addition, the paper discusses the AU-E accuracy and limitations through a diverse and vast array of practical examples including a comparison with thermal modeling and physical measurements.

Jan 1, 2011

By M. Maleki-Moghaddam

At the concentration plant of the Gol-E-Gohar Iron Ore Company (Iran) three 9 m × 2.05 m AG mills (fixed speed) are used in parallel in a dry operation. The performance of AG mills has been lower than the target value indicated by an increase in the product size (P80) from 450 µm (target) to 515 µm and low power draw (50% of nominal). By simulation it was found that increasing the liner lifter face angle from 7 to 30o while keeping the original lifter height (i.e., 22.5 cm) could provide an appropriate charge trajectory. When the proposed liners installed in one of the AG mills, the throughput increased by 27%; when 5% (v/v) balls were added the throughput increased by 40%. The product size also decreased from 515 to 501 µm.

Jan 1, 2012

By David Thomas, Charles Harper, Brian Watters

The Star-Waddy Lakes area lies withjn the northeasterly-trending early- to mjd-Proterozoic volcanoseilimentary-pl utoruc belt referred to as the La Range Domllin. In the area investigated the domajn is composed of the Central Metavolcaruc Belt which is bordered by metasedimentary belts of the Crew Lake and MacLean Lake Belts to the northwest and southeast respectively. The boundaries between the Central Metavolcanic Belt and the flanking belts are generally marked by prominent zones of shearing. The volcanosed imentary succession is intruded by a variety of ultramafic to felsic plutons which, on the basis of lithology, field relations and geochronology, can be broadly subdivided into three groups or episodes. These are: a) an 'early' (ea. 1866 Ma), possibly subvolcanic, episode marked by composite intrusions ranging from gabbro-diorite to granite, b) an 'intermediate' (ea. 1854 Ma) plutonic event distinguished by more homogeneous felsic intru ions ; and c) a 'later' (ea. 1835 Ma) event characterized by leucocratic quartz-rich grani te intrusions. The La Range Domain has undergone polyphase deformation with a strongly developed northeast-trending structural grain. Large scale folding is manjfested by large regional sy nforms, arcuate and triangular patterns associated with many of the plutons and by tight to isoclinal folds within the adjacent metasedimentary belts. Inhomogeneous strain characterizes much of the Star-Waddy Lakes area. With strain gradients being typically very steep, zones of high strain consisting of zones of extreme extensional fabrics, discrete shearing, as well as broader zones of cataclasis and structural slicing, are often juxtaposed with areas of relatively minimal deformation . Dilation structures tend to favour these zones of highly contrasting strain. Late, brittle fractures, faults and shear zones are widespread and appear to be particularly important as loci for gold mineralization. Metan1orphic mineral assemblages indicate that upper greenschist to middle amphibolite facies conditions were attained. Geological and geochenucal evidence indicates that the Central Metavolcanic Belt evolved in an island arc environment. Metavolcanic rocks from both the Waddy and Star Lakes areas display similar characteristics on variation diagrams (e.g. AFM and the Jensen cation plot) in that the dominant trend is calc-alkaline with some basic to intermediate types having a mildly tholeiitic character. A few samples from the southern end of Lower Waddy lake display komatiitic to basaltic-komatiitic afl'injties. The relatively immobile minor and trace elements (e.g. Ti , Zr, Nb, Y, Cr) provide reliable characterization of the metavolcanics, particularly of the basalts, which have highly distinctive trace element properties. Geochemical patterns for the basalts are similar to those for modem volcanic arc basalts. Low initial 87Sr/86Sr ratios (0.7017 - 0.7019) are consistent with a subduction-related origin for the volcanics. The major plutons have chemical properties typical of volcanic arc granites and are largely calc-alkaline in character, showing little chemical variation from the volcanics.

Jan 1, 1986

By Nicholas C. Williams

This paper presents the development of a demand reduction and responsive strategy capable of reducing the electrical power demand, particularly during peak conditions. The work presented includes a case study of a UK colliery. Areas explained are the equipment installation at the mine; the development of the demand reduction strategy including the load profile database and prediction optimisation model. Some live production data from the power meter installation at the mine is also presented.

Aug 1, 2013

By Will Callupe Arzapalo

The success of the minimization of blast induced-damage at Barrick?s Pierina Mine in Peru is the result of an effective working relationship between Mine Operations and the Geotechnical Services Department. A detailed design criteria manual is used to match the blast plan to the site?s varying rock types and slope designs. The Geotechnical Services Department has further developed a comprehensive pre- and post-blast checklist that is critical to the drilling and blasting process as it pertains to wall stability. Wall control blasts are evaluated with quality assurance field checks, vibration monitoring and frequency control, video assisted performance quantification and post excavation face profiling. These detailed observations are compiled and summarized in a blast performance evaluation that is considered to be an ?Industry Best Practice?. The information and detail provided by each performance summary allows accurate blast analysis and effective design refinement.

May 1, 2006

By PAULL. BROUGHTON

Commercial deposits of sodium sulphate are harvested from alkaline lakes that are widespread throughout semi-arid southern Saskatchewan and adjacent southeastern A lberta. Concentration of sodium sulphate occurs in undrained basins under semi-arid conditions. Precipitation runoff from land adjacen1 to 1hese basins leaches 1he mineral from the soils and carries it in solu1ion 10 the basins. The process may be largely subsurface, wherein spring seeps feed basins associated with buried (preg/acial) river valleys. Evapora1ion during summer concentrates salts in the brines, and chilling of water below sa/Uration poin1 during win1er resu//s in deposi1ion of sodium sulpha1e (mirabili1e) crystals on lake bo//oms. Sodium su/pha1e, commercially known as sa/1 cake, ranks second only 10 po1ash in cen1ral-wes1ern Canada's nonme1allic mineral production. // is recovered from su,face brines, directly from intermittenl salt beds, and, until recently, by solution mining of shallow subswface deposits. Sodium sulphate is extrac1ed from seven alkaline lakes and sail f/a1s in Saska1chewan and one in Alberta. Annual produc1ion is approxima1e/y a halfmillion tonnes. Manufacture of kraft paper is the mos1 significanl marke1 for salt cake, but addi1ional quantifies are used in the glass and detergent manufac1uring indus1ries.

Jan 1, 1984

By J G. Brady

This paper discusses the ceramic products used in the construction industry in Canada, as well as the main raw materials used in their manufacture. The products include the! following: (1) clay prorlucts -such as facing brick (glazed and unglazed), common brick, facing tile (glazed and un-glazed), terra cotta, load-bearing tile, non-load-bearing tile, structural clay floor tile, quarry tile, drain tile, sew-er pipe, conduits .and flue liners -which are made mainly from domes-tic common clays and shales and from domestic or imported stoneware clays and low-grade fire clays; (2) floor and wall tiles of the whiteware type, which are produced from ingredients such as kaolin, ball clay, feldspar, nepheline syenite, talc, pvro-phyllite and wollastonite; (3) glass products -the major ones being sheet glass, plate glass and mineral wool (glass wool, and rock or slag wool) -which are made from such materials as silica sand, soda ash, salt cake, feldsspar, nepheline syenite, dolomitic limestone, high-calcium limestone. borax and (for rock and slag wool) various rocks or slags;

Jan 1, 1963

The control of AG mills fed by ores with harder and softer components can be challenging due to the sensitivity of the mill to the ratio of hard to soft components. The mills at LKAB?s iron ore operations are generally well controlled, but there are periods when the throughput falls off dramatically, due to a loss of mill filling. As the ore is upgraded in the sorting plant through magnetic separation prior to milling, this presents an opportunity to manipulate the feed blend in advance. In order to assess this opportunity, a fullscale mill charge was dumped and assayed by size to quantify the build-up of competent waste in the mills. This was supplemented by pilot and locked-cycle laboratory tests at a wide range of feed blends in order to facilitate the development of an appropriate multi-component AG mill model. This paper reports the experimental campaign requirements and methodologies, as well as the major outcomes from analysis and modelling exercises using the industrial, pilot and laboratory multi-component data.

Jan 1, 2012

By N. F. Faria, F. Monteiro, B. O. Taiwo, A. Palangio, D. Fonseca, J. Manuel

This paper explores the interpretation and application of Particle Size Distribution (PSD) curves in optimizing blast design and enhancing downstream operations. The study looks at important factors like D50, D80, fines content, and the uniformity index to show how PSD curves affect the efficiency of different mining processes. The first section of this paper discusses the impact of blast design parameters such as burden, spacing, and explosive energy on the resulting PSD, highlighting their significance for downstream operations like crushing, and material handling. The section concluded that proper alignment of PSD with downstream size classes, like primary crusher inlet sizes, among others, can result in substantial cost reductions and productivity gains. The second part presents a case study from dstgroup aggregate quarry in northern Portugal, utilizing fragmentation analysis to evaluate blast performance and promote continuous improvement. The section specifically examines the effect of stemming material size on blast fragmentation, comparing the impact of using aggregate with a D80 of 12.32 mm versus a D80 of 21.94 mm. The D80 value represents the particle size at which 80% of the stemming material is smaller, making it a critical parameter in evaluating the stemming's effectiveness in confining explosive energy. The drilling and blasting designs were meticulously planned using O-PitDev and O-PitSurface software to ensure accuracy in blast parameters and layout. To assess the fragmentation resulting from these different stemming material sizes, the study utilized WipFrag 4 software. The software allowed for precise measurement of particle size distribution after the blast, as well as a detailed comparison between the two stemming sizes. Results from the analysis indicate that using bigger-sized stemming material led to improved blast confinement, which in turn resulted in better fragmentation compared to the smaller-sized material. The reviewed blast videos demonstrate that the larger aggregate effectively contained the explosive energy within the borehole, thereby reducing premature venting and directing more energy towards breaking the surrounding rock mass. This led to finer fragmentation with D80 = 1240.1 mm, D50 = 685.9, and Dmax = 2580 mm, which is advantageous in downstream operations like crushing and milling. On the other hand, the smaller stemming material (D80 = 21.94 mm) did not confine the explosive energy as efficiently, resulting in coarser fragmentation (D80 = 1327.37 mm, D50 = 750.76, and Dmax = 2370 mm) and reduced overall blast efficiency. These findings highlight the significant influence of stemming material size on blast performance and fragmentation outcomes. 

May 4, 2025