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By Arthur Taggart

A GENERAL awakening of interest among mill men concerning the mechanical efficiencies of their crushing machines is evident from a perusal of the recent files of mining publications. Considering the large part of the power bill which must be debited to the crushing department, such. interest is natural. When, however, the articles on the subject are read, the only statement upon which the writers agree is that fine grinding consumes a large amount of power, and that the machine which, from a concentration standpoint, offends as a slime producer is, for that reason, a serious offender in wasting power. When the mill man attempts to find a method for measuring the power wasted, or to compare the efficiency of one machine with that of another for the same work, he is confronted with a choice of methods which lead to conflicting results. There are two reasons why the work of crushing is not more generally understood. In the first place, experimental data are difficult to obtain; and in the second place, before these data can be put in general form certain assumptions must be made. A difference in these assumptions by different investigators has led to conflicting conclusions. This conflict has resulted in the proposal of two laws. The first of these is the so-called Rittinger's law, which is variously stated as follows: . "The work of crushing is proportional to the reduction in diameter."-R. H. Richards: Ore Dressing, p. 304 (1903). "The Work required to crush rock is very nearly proportional to the reciprocals of .the diameters crushed to."-A. Del Mar: Engineering and Mining Journal, vol. xciv, No. 24, p. 1129 (Dec. 14, 1912). "The Work done in crushing is proportional to the surface exposed by the operation, or better expressed for this purpose, the Work done on a given mass of rock is proportional to the reciprocal of the diameter of the final product, assuming that all the mass has been reduced to one exact size, which is only theoretically possible."- A. 0. Gates: Engineering and Mining Journal, vol. xcv, No. 21, p. 1039 (May 24, 1913).

Jan 1, 1914

By H. E. Bartlett

Previously, the grade of ore at the shaft head was taken to be equivalent to the grade measured in the faces by means of chip sampling. The tonnage mined from the stopes is determined from survey volumetric measurements. Estimates of shaft head grade and tons include the grades and tons of waste, development ore, sweepings from old areas, and other sources, not all of which are sampled. The tonnage of ore delivered at the shaft is often estimated using skip factors. These factors are subject to manipulation and not all the skips are filled to the same level. Consequently, both the grade and tonnage of ore delivered at shaft head are subject to a measure of uncertainty. For many years, these uncertainties have led to disputes between shaft managers and metallurgists over the reconciliation between shaft tons and grade and plant tons and grade. These problems are compounded when several shafts feed a central processing plant. Shaft bonuses and a shaft?s profitability are affected by poor gold allocation methods. Without measurement of the grade and tonnage from all shafts, a poorly performing shaft could benefit at the expense of the other shafts. In order to improve the measurement of grades and tonnage at the shaft head, which is the custody transfer point between the mine and surface operations, South African mining companies have developed a methodology to measure the mass of ore from each shaft using rail weighing equipment, and the run-of-mine grades delivered from each shaft using cross-stream (go-belt) sampling. Harmony Gold Mining Company has implemented the rail weighing and go-belt sampling methods for their shafts. This paper deals with the statistical tests on the procedures and control measures that are in place to establish confidence in the gold allocations. The metal accounting system for examining the allocations follows the first principle of the AMIRA P754 Metal Accounting Code of Practice namely ?The metal accounting system must be based on accurate measurements of mass and metal content. It must be based on a full check-in check-out system?. The check-in is mass and grade of run-of-mine ore and the check-outs are the plant input grades and the recovery and residue from the plant. The balances between check-in and check-out are the subject of the statistical analyses, which have been conducted using the classical statistical tests and procedures that are generally used in all fields of endeavour. It is concluded that the gold allocation procedures that rely on go-belt sampling and rail weighing of ore from shafts has been successfully implemented, and that the system, which has been in operation at Harmony since 1993, has won general acceptance from stakeholders, with the gold allocations being perceived as being fair to all. Apart from the on-mine issues of bonuses and fair allocations, accurate metal accounting assists in the broader issues of generating accurate production figures for management. An added benefit of weighing and sampling ore at the shaft head has been that shafts now have daily moving average grades, which are used for grade control. The go-belt sample at the shaft head is the first accurate sample of production in the value chain from face sampling to gold bullion. Keywords metal accounting, go-belt sampling, run-of-mine sampling, codes of practice, AMIRA P754, train weighing, statistical analysis, gold split, allocation of gold, metal balance, grade control.

Jan 1, 2014

By P Dare-Bryan

Many mines have an interest in improving fragmentation at their sublevel caving operations in order to produce better ore recovery, lower dilution and less overbreak. The primary requirement for the extraction phase is to understand and optimise the flow of the fragmented rock in the rings including interactions with the caved material. Rock fragment formation is critically controlled by the blasting phase and can potentially be improved by optimising the dynamic energy distribution throughout the ring through the crafting of explosive charging, priming position and initiation timing. The modelling of fragmentation and damage due to explosive loading of complete rings is inherently and irrevocably a three dimensional problem that is difficult to address. We have therefore initially utilised an analytical code to predict vector peak particle velocities and energies that are restricted to the plane of a ring. Dynamic energy distributions based on common timing and priming practice have been compared with other possibilities using conventional pyrotechnic detonators as well as initiation designs based on accurate electronic detonators. The energy distribution can be quite subtly tailored to increase energy in appropriate regions or to decrease energy in the brow region for example to potentially reduce instability and the probability of misfires. We have also recently developed the capability to efficiently model particle flow using four-sided polygonal elements. Despite the 2D limitations of such a model this results in more realistic fragment interaction and stacking than can be achieved using conventional rigid circular elements. Modelling of a typical sublevel caving ring utilising these elements activated by the correct dynamics offers considerable potential for understanding the influence of the blasting energetics and timing on the complex flow and extraction process.

Jan 1, 2005

By G. Girard, A. Larouche, J. Colbert, G. Doyer

"The Direct Chill (DC) casting process uses water to extract heat from molten aluminium and helps to control the solidification process. Many factors such as the water flowrate, water distribution around the mould and water impingement angle have significant impacts on the solidification process. It has been noted that the chemistry variation of the cooling water is also an important factor that has a major impact on the presence of defects on the ingot. A significant change in the water quality can produce defects on the ingots such as cracks, bleedouts, or runouts. Understanding the impact of water quality variation on ingots is therefore essential to ensure good ingot production and reduce defects. This paper aims at describing a new approach to follow the variations of process water quality over time. It is an original and innovative method, which is still under progress. A water quality and stability index (WQSI) is under development to evaluate changes in the water composition based on water physico-chemical parameters. The index is calculated by comparing each of the actual parameters with historical values previously modeled by a statistical law. It is a dimensionless number having an adjustable weight factor for each parameter. The positive or negative contribution of each parameter, with respect to the water cooling capacity, is taken into account to calculate the WQSI. The current selection of parameters is based on knowledge from the literature as well as on empirical experience. When the developed tool will be mature, it could be used to track water variations upon time and anticipate problems, enabling proactive actions to ensure good ingot production.INTRODUCTIONFor many years, the impact of water chemistry on heat transfer, during DC casting, has been of great interest in the aluminum industry. Several papers have been published to demonstrate the correlation between water chemistry and heat transfer (Yu, 1985), (Gildemeister, 2014). Although numerous studies do present the effects of water parameters (i. e. hardness, conductivity, oil, and grease…) on cooling capacity under controlled experimental conditions, they do not reflect the strong interactions between the parameters and variations present in a casting center. Wells and Cockcroft also demonstrated that laboratory based studies are not fully representative of real heat transfer in an industrial casthouse, there are too many factors involved (Wells, Li, & Cockcroft, 2001). In this context, it becomes very difficult to obtain a quantifiable relationship between the capacity of water to cool the ingot and ultimately the performance of the cast."

Jan 1, 2016

By J Brisebois

The derelict Lake George Mine is located at Captains Flat in NSW, acommunity of approximately 450, which lies some 43 km southeast ofQueanbeyan. It straddles the Molonglo River, which flows toQueanbeyan and thence the nationÆs capital, Canberra. Catastrophictailings dam failures in 1939 and again in 1943 resulted in the transportand dispersion of thousands of tonnes of sulfidic wastes downstream,with terminal settling in Lake Burley Griffin, in Canberra. The mine wasclosed in 1962, and major rehabilitation in 1996 - 1997 addressed the riskof further collapses, and reduced water flow into the underground mine.Nevertheless, acid mine drainage from the mine, and seepages ofcontaminated water from the capped tailings dumps continue to pollutethe Molonglo River for some 30 km downstream.Environmentally, the highest contributor to the total contaminant loadof the river is the Main Adit Spring. This was shown to be a constantgroundwater-driven acidic flow, with contaminant generation resultingalmost entirely from the ingress of gaseous oxygen into the undergroundworkings. Long term remediation may be possible at source, by pumpingreducing gases into the sealed mine, but the probability of thistechnology succeeding to the required degree is yet to be demonstrated. For at least the medium term, treatment of contaminated water flow(s)will be required to achieve any of the water quality targets that may beapplicable to the river. Three methodologies were short listed, but anumber of strategic uncertainties remain, which the NSW Derelict MinesProgram will address. Future risk arising from long-term contaminantrelease by the treatment wastes must be considered. Appropriate data,and realistic long-term performance parameters regarding the relativestabilities of water treatment sludges are scarce, and make properevaluation of competing technologies problematic.

Jan 1, 2003

By B. W. Gandrud

In 1927 Coghill5 showed that there is considerable liberation of the iron oxide mineral from gangue when the Alabama red iron ores are ground to a size suitable for classification and tabling, but that there is not sufficient to make possible a high extraction and a high grade of concentrate. Coghill states: High extraction and high grade of concentrate cannot be obtained without grinding to a size comparable with flotation or cyanide practice. If grade of concentrate be sacrificed the grinding need not be so severe.

Jan 1, 1934

By Kenneth B. Ray, George R. Gardner

The application of wet cleaning processes for the beneficiation of bituminous coal has created in some localities a problem in the recovery and disposal of fine solids in the washery water. The maximum allowable concentration of these solids in the washery circuit that is consistent with satisfactory operation of the cleaning units remains a controversial point. It has, however, been generally recognized for a number of years that some degree of control should be exerted over the concentration of such material. Early methods of control involved intermittent or continuous discharge of a portion of the plant slurry into near-by streams or settling ponds, and this persisted even after the adoption of settling tanks or cones. The practice is not always economical, for two reasons: first, because of the waste of coal, and, second, because it is sometimes difficult and expensive to obtain an adequate supply of water. The latter condition often is aggravated by chemical treatment of the make-up water that is necessary to prevent excessive corrosion of plant equipment. The modern approach to this problem has been the adoption of continuous thickening devices. However, because such equipment occupies considerable plant space and is somewhat expensive to construct, the practice of loading these thickeners beyond their rated capacity is widespread, therefore often the performance of such units is not altogether satisfactory. In recent years, European washeries have increased the rate of settling of the fine coal from the slurry by the addition of specially prepared starch solutions. These reagents cause flocculation of the solid particles and thereby increase the rate of sedimentation. The first important work on the flocculation and clarification of coal slurry was done by Henry, who developed the Henry process and used it successfully in Belgian preparation plants. Soon after the initiation in Belgium, the process was introduced into England, where subsequent investigation developed similar processes, which were successfully used in

Jan 1, 1940

By George R. Gardner, Kenneth B. Ray

The application of wet cleaning processes for the beneficiation of bituminous coal has created in some localities a problem in the recovery and disposal of fine solids in the washery water. The maximum allowable concentration of these solids in the washery circuit that is consistent with satisfactory operation of the cleaning units remains a controversial point. It has, however, been generally recognized for a number of years that some degree of control should be exerted over the concentration of such material. Early methods of control involved intermittent or continuous discharge of a portion of the plant slurry into near-by streams or settling ponds, and this persisted even after the adoption of settling tanks or cones. The practice is not always economical, for two reasons: first, because of the waste of coal, and, second, because it is sometimes difficult and expensive to obtain an adequate supply of water. The latter condition often is aggravated by chemical treatment of the make-up water that is necessary to prevent excessive corrosion of plant equipment. The modern approach to this problem has been the adoption of continuous thickening devices. However, because such equipment occupies considerable plant space and is somewhat expensive to construct, the practice of loading these thickeners beyond their rated capacity is widespread, therefore often the performance of such units is not altogether satisfactory. In recent years, European washeries have increased the rate of settling of the fine coal from the slurry by the addition of specially prepared starch solutions. These reagents cause flocculation of the solid particles and thereby increase the rate of sedimentation. The first important work on the flocculation and clarification of coal slurry was done by Henry, who developed the Henry process and used it successfully in Belgian preparation plants. Soon after the initiation in Belgium, the process was introduced into England, where subsequent investigation developed similar processes, which were successfully used in

Jan 1, 1940

By C. V. O?Keefe

Volumetric flow and mass flow are two of the critical process parameters needed to understand and control flotation processes, with others being levels, densities, and metal analyses. Both volumetric flow and density measurements can be adversely influenced by entrained air, which will result in artificially high flow readings and artificially low mass flow calculations. Array-based instrumentation has been used successfully to accurately measure both volumetric flow of aerated slurry and the quantity of air entrained within the slurry. The latter measurement has been used to correct density measurements, thus providing an accurate mass flow calculation. A case study involving the use of this technology to measure both flow rates and air entrainment levels at Eland Platinum processing facility (concentrator) will be covered in detail. In addition, with certain traditional flow meters such as electromagnetic meters, erroneous volumetric flow rates can be recorded in the presence of paramagnetic or ferromagnetic material. The use of this array technology to avoid spikes in measurements due to steel fragments passing through the pipe will also be discussed.

Jan 1, 2012

By A. K. Mukherjee, J. E. Dorn

A modified Peierls model is Presented for the rationalization of the asymmetric plastic behavior of single crystals of bcc metals. The model assumes that the asymmetry of slip arises from a tendency toward asymnetric splitting of the a/2[111] screw dislocation. The approach differs from former theories however insofar as it assumes that the splitting is so minor ttut only the core energy is modified and that no real partial dislocations exist. As a consequence, the concepts of a line energy of a partial dislocation, interaction between partial dislocations, constriction energies, and stacking fault energies are not invoked. Instead the present analysis is based on a specialization of the Peierls mechanism which assumes an asymmetric dislocation core which is modified by the effective stress and permits cross slip. The theoretical predictions on the stress-temperature -strain rate and orientation relationship agree qualitatively with the limited experimental data that are currently available. Additional experimental work over wider ranges of conditions is suggested to provide more complete evaluation of the possible validity of this theory. RECENT experimental investigations on the low-temperature thermally activated deformation of bcc metals have confirmed the existence of several unique orientation effects. These will be described with the aid of the standard stereographic projection shown in Fig. 1. If slip were determined exclusively by crystal symmetry, the deformation characteristics would be symmetrical about X; = 0 and independent of whether the specimen were subjected to a tension or a compression stress, regardless of what combinations of slip planes might be operative. Experiments, however, reveal that neither of these symmetry based predictions are valid. A schematic representation of the commonly observed asymmetry of both the yield stress and the plane of slip is shown in Fig. 2. A reversal of the direction of stressing results in an inversion of the asvmmetric behavior. 1) Asymmetry bf Yielding. A detailed investigation on the asymmetric yielding in an Fe-2.7 pct Si alloy was repoEted by Tabka, Tikeuchi, and Furubayashi,1 Similar trends have been observed by Keh and Nakada on iron,' Stein on molybdenum,3 Bowen, Christian, and Taylor4 on columbium, Sherwood, Guiu, Kim, and pratt5 on Ta, Cb, and Mo. The degree of asymmetry not only depends on the atomic number and alloying but, in all cases that have been examined, it also increases with a decrease in the test temperature. 2) Asymmetry of the Slip Plane. Asymmetric $I-~: curves of the type shown in Fig. 2 have been found for several Fe-Si alloys6-8,1 and columbium.4 There is general agre_ement that near X: = 30 deg slip occurs by the [111](211) mode so that ?1 = 30 deg. As X; decreases it appears that [111](211) to [111](101) cross slip takes place leading to slip band traces on intermediate planes. And as xi approaches 0 deg, sharp

Jan 1, 1970

By Levi Guzmán Rivera

Dada que la cantidad de energía utilizada en el proceso de conminución equivale al 3% de la energía mundial, es necesario buscar un proceso de mejora y no limitarse a la ineficiencia de las tecnologías convencionales de molienda SAG y bolas puesto que utilizan para el proceso de molienda entre 3 a 5% del total de la energía consumida. En ese sentido, el presente texto señala la optimización del consumo de energía en dicho proceso, de forma que mediante el uso de métodos discretos, se pueda determinar la óptima trayectoria de la carga moledora y así mejorar la eficiencia de la molienda.

Nov 1, 2012

By M. B. Mirza, F. D. Wright

Models of photoelastic material were made to simulate a mine roof that had cracked over the edge of the pillars and at the center of the span. Models were restrained from moving laterally outward so they acted as flat arches. Uniform loading was approximated by loading the models along the bottom at eight points spaced equally along the span. It was found that for models with depth to span ratios of 1:3 to 1:6, the maximum stress at the top of the center crack could be calculated directly by the following formula: Elastic stresses in mine roofs composed of sedimentary beds are usually computed by assuming that the immediate roof bed acts as a fully restrained beam uniformly loaded by its own weight and by overlying strata. Very little work has been done to analyze stresses in roof beds or beams which contain vertical (normal to the span) or nearly vertical cracks. Yet every mining engineer has walked safely under roofs that show such cracks. As far back as 1933, Bucky' made model studies of the effect of vertical cracks on the behavior of horizontal strata. He did not make any analysis of the stresses but did show that a cracked beam will stand safely under certain conditions, and also showed that a beam restrained at the ends can crack over the supports and in the middle without collapsing as long as the ends of the beam are not free to move laterally outward. The beam, or rather, part of the beam, acts like a flat arch and supports itself by pushing outward at its supports. In 1941 Evans2 treated analytically the problem of a beam cracked along so many planes perpendicular to the axis of the beam that no tensile stress can operate. He named this type of beam a "voussoir-beam" by analogy to the voussoir-arch. In 1962 ~a~cocks~ modified Evans's solution by including the effect of the shear stress component of the resultant force at the abutment. This paper gives the results of a photoelastic study of the stresses in a beam containing vertical cracks in the center of the span and over the supports. A

Jan 1, 1963

By Kent E. McElhattan

The selection of efficiency machinery for an underground mining system is a highly skilled art. Modern techniques of mining, based on sound engineering principles, increase the probability of success. These techniques alone, however, will not guarantee maximum productivity. Human judgement, based on experience with mineral seams and geological conditions, remains the single most important input of the decision-making techniques in selecting high production equipment. The myth of a "Scientific Selection Factor" is absurd. The belief that such a factor exists can waste the time and resources of mining management. Until now, mineral extraction from the original desposit has defied all attempts to systematize the knowledge. Computerized programs are excellent for solving underground transportation and ventilation problems. The variables can be defined by mathe¬matical techniques in this type of problem. By the same, token, "Trial and Error" is not an acceptable method in today's technology-minded world. The candid truth of the matter is that practically nothing is known of the phenomenon of extracting a mineral from the solid. Likewise, local conditions make it nearly impossible to predict accurately the reaction of surrounding strata after the mineral has been removed. The unpredictability of the environment resulting from extracting minerals underground will remain a challenge to mining engineers for many years. In fact, we are barely beyond the trial and error stage at this point in time.

Jan 1, 1969

In taking the chair at this meeting of the Australasian Institute of Mining Engineers, I feel it my duty to thank you for the honour you have conferred upon me by electing me President for the current year, especially as Bwken Hill is the birth place of the Institute and myself one of its foundation members. I will now endeavour to entertain you for a short time with a few reminiscences of Broken Hill from its inception. Beyond the fact that the mines on the Broken Hill lode have paid to their shareholders in dividends and bonuses no less a sum than £896,500, created and maintained a city of 25,000 inhabitants. supplied a large proportion of the revenue on the railways of a sister state, and made their influence felt throughout the 'commercial centres of the world, very little, I regret to say, is known by the general public as to the form, extent, or vastness of the ore bodies, or the mode of occurrence of the various classes of ore which have been found to exist in them. It has occurred to me that a short paper by the senior mining manager on the line of lode might prove interesting. There has been great difficulty in getting some of the figures relating to the field, the total output of metals from the Broken Hill mines not being obtainable from the Mining Department in Sydney j but an inscription on the foundation stone of the Technical College recently opened in this City, sets forth that the Broken Hill Mines started in 1880...

Jan 1, 1903

The Namakwa Sands heavy mineral deposit is located along the West Coast of South Africa and the mine is a world class producer of high quality zircon, ilmenite and rutile concentrates from essentially unconsolidated marine and aeolian sands of Cainozoic age. The objective of this study was to characterize rutile with respect to distribution, grain size, textures, colour and mineral chemistry within the Namakwa Sands orebody with the aim to explain its overall poor recovery. A representative suite of heavy mineral concentrates from various sections of the orebody and stages in the recovery circuit has been investigated microscopically and by means of SEM-EDS and LA-ICP-MS. The rutile grain size distribution displays a wide range and is directly related to its variable chemistry and consequently density. In placer deposits such as Namakwa Sands, mineral sorting is a function of hydraulic equivalence, and high density grains are smaller than the lighter grains when deposited under similar conditions. Grain size is also a function of sediment maturity, and the highly mature red aeolians and (RAS) component of the deposit has a coarser grain size than the less mature orange feldspathic marine sands (OFS). The above primary characteristics of rutile are responsible for the loss of the coarse-grained fraction during screening and, given the empirically determined relationship that increased substitution elements reduce conductivity, also during electrostatic separation. Provenance studies using geothermometry have shown that the heavy mineral suite has been sourced mainly by the proximal medium-to high-grade Namaqualand Metamorphic Complex. The positive relationship between substitution elements and temperature of formation explains their high concentration in rutile of this deposit. The heterogeneity of rutile, produced by the combination of a high temperature primary source and the typical marine-aeolian placer genesis is deposit specific and unfortunately not conducive to high recovery levels in the current Namakwa Sands beneficiation circuit.

Jan 1, 2010

By R. V. Ramani, J. M. Mutmansky, J. Qin, J. Organiscak, R. Bhaskar

"The results of a two—year research study into the relationship between quartz dust levels in host material and in respirable dust in continuous mining sections are presented in this paper. The experimental design and sampling procedures for underground data collection are outlined. The sample analysis procedures for quartz determination, the ¬analysis results and conclusions of the study are discussed.The results of the study indicate that the quartz contents of the airborne dusts are generally higher than those in the host material and bulk samples; and the quartz content increases with decreasing size of the airborne dust. The source for quartz appears to be the roof and the floor materials rather than the coal seam itself. It also appears that the material containing quartz may be breaking into a different size distribution than coal.INTRODUCTION Without doubt, the potential for permanent and severe health damage exists due to respirable dust in the work environment. The attention directed at the reduction of respirable airborne dust to levels below those specified in the law and the achievements to date in meeting these standards are but a reflection of the recognition by many parties government, industry, and labor - of this potential and the need to reduce it.The ambient dust standard in underground U.S. coal mines (2 mg/s3) has a qualifier (Code of Federal Regulations, Title 30, Part 71.101) to determine an applicable dust standard when a respirable dust sample contains more than 5% quartz. This standard is frequently referred to as the ""reduced standard."" The reasoning behind the reduced standard is the relative toxicity of the quartz dust compared to that of pure coal dust. As the quartz levels in samples increase, the applicable standards become more stringent and maintaining the workings in compliance becomes increasingly difficult.The number of underground work locations operating under the reduced standards has shown a dramatic increase since 1980 (Figure 1). For comparison purposes, there were no mining entities operating on reduced standards prior to 1977. The emergence of this problem, which affects the health and safety standards on one hand and creates severe compliance problems on the other, has been rapid."

Jan 1, 1990

By Walter Herd

For the past eight years, No. 2 mine of the Cumberland Railway & Coal Co., Springhill-a subsidiary of the Dominion Coal Co., Ltd.-has had an unenviable reputation for bumps. As the workings extended in certain directions under a greater cover, these bumps increased in severity and it became apparent some change in the system of extraction was necessary if the mine was to continue to operate. In September, 1924, following several fatalities resulting from bumps, the Mines Depart~ of Nova Scotia requested from the U. S. Bureau of Mines the services of George S. Rice, Chief Mining Engineer of the Bureau, to report on the conditions then prevalent in No. 2 mine, Springhill. °This request was promptly and courteously granted. Mr. Rice made a full investigation and a comprehensive and helpful report, which is embodied in the 1924 Annual Report of the Department of Public Works and Mines, Nova Scotia. Prior to 1925, the room-and-pillar method was employed, and in January of that year (there being a cessation of work from March to August due to labour troubles) longwall retreating on the west side, on a limited scale, and in a small section on the east side of the mine, was started. The major portion of the east side was not operated on this system till August, 1926, as there was developed a considerable area ready for pillar extraction between an arbitrary boundary and a point up to which bumps of any consequence had not been experienced. In ?the interval, development work for longwall retreating below the 5,900-foot level was vigorously pushed.

Jan 1, 1929

By L. A. Heim, E. A. McCullough, N. J. Amiotte

The explosives industry is subject to the laws of multiple jurisdictions, which means that several regulatory agencies are responsible for governing different aspects of the commercial explosives industry. Among these regulators are the Mine Safety and Health Administration (MSHA) who implements codes that better ensure both occupational and public safety which is influenced by professional blasting projects and the Bureau of Alcohol, Tobacco, Firearms and Explosives (ATF), who enforces commercial explosives storage and security, among other responsibilities.

Jan 1, 2014

By A G. Grice

Surface disposal of mining processing wastes is usually perceived as an undesirable consequence of meeting societyÆs needs for metals and minerals. Environmental disasters associated with the discharge or storage of such wastes, though infrequent, can have very damaging consequences for the environment and the industry. Reducing or eliminating the potential for such events may be achieved by placing some or all of these processing wastes back into the voids created by the mining operations. Processing operations may limit or enhance the opportunities for using tailings and other waste products as a backfill material. Backfilling is a truly multi-disciplinary process and requires the informed involvement of mining, metallurgical, engineering and environmental specialists. This paper describes the range of fill types currently being used in Australian mines, particularly those utilising processing wastes as constituents. It focuses on how processing operations can affect the suitability of these materials for the various types of fill, the resulting fill properties, fill performance and costs.

Jan 1, 2002

By J. J. Crowe, G. K. Burgess

PAGE, Theories of Allotropy of Iron........... 667 Previous Determinations of A2 and A3 in Iron........ 668 Critical Ranges as Determined by Expansion........ 669 Critical Ranges by Thermoelectric Observations........ Critical Ranges in Terms of Crystalline Structure....... 670 Critical Ranges as Detected by Mechanical Methods...... 670 Critical Ranges as Given by Electrical Resistance....... 670 Critical Ranges as Determined Magnetically........ 671 Critical Ranges as Determined Calorimetrically....... 673 Critical Ranges from Heating and Cooling Curves....... 673 Summary of Previous Determinations of A3 and A2 in Pure Iron.... 675 The Present Investigation............ 678 The Experimental Method........... 678 Arrangement of the Apparatus.......... 679 The Measurements of Temperaturc.......... 679 The Measurement of Time........... 679 The Thermocouples............ 681 Temperature Scale and Calibration of Thermocouples...... 681 Mounting of Thermocouples........... 682 The Differential Measurements.......... 682 The Furnace and Accessories........... 682 The Vacuum Pump............ 683 The Energy Supply............683 The Rheostat.............. 683 The Materials Investigated............ 685 Sources of Iron............. 685 Preparation and Description of Samples......... 685 Chemical Analyses............. 687 The Experimental Data............. 688 The Preliminary Observations.......... 688 The Definite Series............691 Description of the Plotted Curves........... 694 The Effect of Plotting at Different Intervals........ 694 The Curves of the Definite Series.......... 695 . The Curves of the Preliminary Series......... The Existence of Other Critical Ranges......... 696 The Microscopical Examination........... 697 Summary and Conclusions............ 698 Bibliography............... 701 Plates................ 704

Jan 1, 1914