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By A. D. Mutch

IT has long been known that there are wide variations in the thermo-electric potential of pyrite. In the last few years a study of this variation and other properties of pyrite was carried out at the University of Toronto by F. G. Smith on specimens of pyrite from many localities.1-3 As a result of these studies the pyrite geothermometer was developed by Professor Smith. The calibration of the pyrite geothermometer is based on the assumption that "crystals of any one electronically conducting mineral species deposited at a high temperature have a more positive thermo-electric potential than crystals deposited at a low temperature, and the thermo-electric potential varies continuously between any given limits of the temperature of deposition." The cause of the variation was postulated to be crystal discontinuities, such as lineage boundaries. As yet this postulated relationship has not been established as fact. On theoretical grounds it is indicated that the thermo-electric potential of pyrite or any other partial conductor is dependent upon a number of factors as well as the temperature of formation. For this and other reasons there is ground for doubt as to the validity of the relationship postulated above. To date a large number of readings have been made with the pyrite geothermometer which have been recorded in terms of the calibration of the instrument in degrees centigrade. So far most of the pyrite tested by this technique has been in association with gold ore. The values for the temperature of formation obtained in these tests appear to be reasonable and, in relation to geological problems, to vary in the right sense. Determinations on some specimens not in association with gold ore have given readings that appear to be much too high, taking into account the probable maximum temperature of the host rock. To sum up, although many determinations made with the. pyrite geothermometer appear to give reasonable values for the temperature of formation of the pyrite, there are at the same time other readings which seem anomalous. On the basis of an examination of available results of work with the pyrite geothermometer, two problems must be considered: 1-the validity of the calibration of the pyrite geothermometer; and 2-the nature of the distribution of the thermo-electric potentials of pyrite in relation to the specific problems that have been examined. In this paper the academic problem relating to the validity of the calibration of the pyrite geothermometer will not be discussed. Instead, the variations of thermo-electric potential of pyrite in association with gold veins will be discussed with regard to specific examinations. The first of these two examinations deals with the distribution of thermo-electric potentials of pyrite in association with a gold vein and was carried out in the McIntyre Porcupine Mine, Schumacher, Ont., in the summer of 1949 as part of a study with this technique. The second problem deals with variations of thermo-electric potential of pyrite in gold ore specimens from all over the world. The specimens used in this study were part of the private collection of Frank Ebbutt, geologist for Howe Sound Exploration Co. Results of Detailed and Reconnaissance Studies In discussion of the results of determinations made in these two studies the values of thermo-electric potential in millivolts will be used. It will be noted that in some of the diagrams the determinations are recorded in terms of pyrite geothermometer temperatures. Fig. 1 is a calibration chart relating, pyrite geothermometer temperatures and thermo-electric potentials. A full discussion of the technique is given in the original paper by F. G. Smith.' Briefly, the determinations are made by placing two stainless steel probes, one hot and the other cold, against the surface of the pyrite crystal being tested. The thermo-electric potential that is developed is read with a null-point potentiometer that is calibrated in what

Jan 1, 1952

By W. P. Campbell

THE investigation herein described had its beginning in the chemical laboratory of the Department of the Interior, Calgary. After the transfer of the natural resources to the Province of Alberta in 1930, the laboratory was closed and the investigation discontinued, but in 1933, with the permission of the Director of the Mines Branch, the work was again taken up and brought to completion. OBJECT OF THE INVESTIGATION The object of the investigation as originally planned may be stated as follows : (1) To determine the chemical composition of the limestone formation underlying Turner Valley gas field. (2) To identify 'beds' in the limestone, if such existed, with definitecharacteristics which would mark them from immediately overlying and underlying measures. (3) To relate information acquired on the composition of the limestone to the occurrence of oil and gas in the particular section examined. Definite information was lacking as to the exact chemical composition of the limestone and of its variations in composition at various depths. The formation was generally referred to in the field as the 'dolomitic' limestone, but the depth to which the dolomitized zone extended and the degree to which the limestone generally had become dolomitized were not known. Jn the drilling and logging of wells, the recognition of suitable 'markers' in the various formations is a matter of some importance. Markers are generally relatively small but widespread members of a formation, which have some characteristic property, of colour, composition, or form, by which they may be identified, and with them, consequently, the formation of which they form a part. In the formations above the limestone in Turner Valley, a number of well defined and easily recognized markers are known to geologists, drillers, and others interested in the field. Many of these markers Hand out in sharp contrast to adjoining measures and are therefore easily recognized; in others, the contrast is less sharp, with consequent difficulty in identification.

Jan 1, 1937

By Charles E. Locke

CONTINUED expansion of gold mining in 1935 led to further developments in treatment methods. In base metals and non-metallics progress is also noted, coincident with greater activity. Statistics are not available yet for 1935 but the trend is shown by 1934 when nonferrous ore mined in the United States was 35,840,737 tons, or 45 per cent more than in 1933. The amount concentrated was 27,167,530 tons or 75.80 per cent of the ore mined, and flotation was used in whole or part on over 96 per cent of the ore subjected to concentration. In the crushing field a new model of the Fahrenwald crusher is promised soon. The Symons fine-reduction cone crusher and Traylor secondary crusher continue to gain as intermediate crushers. A new book entitled "Crushers for Stone and Ore" by William T. W. Miller appeared. The Hadsel mill has been improved still further and a number of installations are in satisfactory operation showing economical power consumption and low wear. An attempt in a

Jan 1, 1936

By O Bøckman, E Thisted, K H. Esbensen, U Thisted

Process industry measurement systems, once installed, are often interpreted as static black boxes delivering correct values as long as not proven otherwise. After first implementation and set-up with a distinct sampling frequency, they often run ‘forever’ without checking if this frequency is still appropriate with respect to changed raw materials characteristics, different process conditions or changed customer specifications. Failure detection is based on operator experience and skills (often close to an art form) and/or from control measurements of reference samples. This may well be too late for many industrial process contexts. An appropriate sampling frequency is a critical success factor for valid monitoring and control.Based on the methodology of variographic analysis we here illustrate how to design/re-design/optimise an industrial measurement system in terms of sampling frequency, discretisation and accuracy based on customer specifications. We also show how variographics can be used to monitor for deviations. We use experience from process industry to show how these methods can be used not only for product characterisation, but also for monitoring input factors, process indicators and measurement of these. The information hidden in the vast amounts of data that is continuously created in modern process industries can be utilised to cut costs in the measurement activities and at the same time contribute to early error detection. We illustrate by case examples from production processes at a nickel refinery, supporting the Lean and Six Sigma methodology. This will provide a standard operating procedure for monitoring the stability of measurement systems, and provide a basis with which to discuss how a variographic approach can be integrated with the Lean and Six Sigma approach as an online tool to map, characterise and validate measurement systems and thus avoid more cumbersome measurement system analysis in full detail.CITATION:Thisted, E, Thisted, U, Bøckman, O and Esbensen, K H, 2017. Variographic case study for designing, monitoring and optimising industrial measurement systems – the missing link in Lean and Six Sigma, in Proceedings Eighth World Conference on Sampling and Blending , pp 359–366 (The Australasian Institute of Mining and Metallurgy: Melbourne).

May 9, 2017

By John A. Gann

THE Iron and Steel Division has shared in the general increased activity of the industry. The fall meeting at Chicago was not only well attended but particularly characterized by the virility and scope of the discussions at the various sessions. The program for the February meeting indicates a similar interest in technological developments. The Blast-Furnace Committee continues to have the hearty support of the industry with discussions of such a nature as to be of real help to the operator as well ass the technologist. The work of the Open-Hearth Committee continues to be outstanding, with ever increasing attendance and ever mounting interest in the meetings. Formation of a conference on the physical chemistry of steelmaking has given those interested in this phase of work an organization allowing the same

Jan 1, 1936

By Chris Lynagh, Collin Schroeder

Determining the best excavation method is a critical step in successful tunnel construction. In rock tunneling, several factors can dictate the excavation method, such as: (1) rock characteristics, (2) tunnel size and length, (3) nearby utilities/infrastructure, and (4) contractor’s preferred means and methods. This paper discusses the various rock tunnel excavation methods utilized on the Doan Valley Storage Tunnel Project (DVT), a project constructed as part of the broader Project Clean Lake in Cleveland, Ohio. The differences between each method’s initial and final lining, geology, and cycle sequence will be analyzed and compared.

Jun 13, 2023

By Louis W. Cope

In the early 1900s, vat leaching was a common process for treating gold ores. It was, however, labor intensive and was abandoned in the 1930s. Technology advances in recent years, though, have caused vat leaching to once again become a viable alternative in precious metal processing. At least three plants have been built. Others are being considered as alternatives to agitation and heap leaching. What is vat leaching? Vat leaching is a process in which ore is crushed and screened to a size of between 6 to 12 mm ( 0.25 to 0.5 in.) This crushed ore is placed in a large, boxlike container where an alkaline-cyanide solution is added and later removed. The process might be likened to indoor, controlled heap leaching. The ore does not require fine grinding in a ball mill. Therefore, capital and operating costs are lower than agitation treatment. But the extra handling causes costs to be slightly higher than heap leaching. If any amount of fines is present, the ore should be agglomerated.

Jan 1, 1991

By K. Zabrusky

Epithermal vein systems are often modeled as simple, discrete high grade structures within low grade host rock. However, geologic setting and structural preparation influence how such vein systems form. In structurally complex areas, veins may follow several structural trends and occur on a variety of scales. Geologic and grade modeling within such complex structural regimes can be difficult as individual veins cannot be delineated from host rock based on grade alone. Structurally complex areas require a different approach than traditional discrete vein modeling. Presented herein is a study of a structurally complex area that hosts vein mineralization to be mined underground. All veins could not be modeled discretely, so an alternate approach was needed. Major structures were modeled using surface and underground mapping and were then used to create a structural trend for the area. This trend was used to guide 3D grade shell construction. The grade shells replaced 3D vein solids for grade estimation in order to mitigate the influence of high grade samples. Structurally guided grade shell construction offers a more effective modeling method than discrete vein modeling in deposits that are structurally complex.

Feb 23, 2014

Declines for both access and rock haulage are replacing shafts for mining relatively shallow deposits of down to 1000 m below surface. Where the orebody does not outcrop and where surface access may be restricted, it is often necessary to drive long dead end headings before a connection can be made back to surface and a ventilation circuit created. In hot tropical areas and where large diesel powered equipment is used for mining, heat may also be a problem at these relatively shallow depths. The normal design criteria based on statutory limits and good practice used to evaluate ventilation and cooling requirements in terms of air velocity, dust, gases, diesel exhaust dilution, re-entry after blasting and heat stress are described in this paper. These are then used to design and cost alternative auxiliary ventilation systems that include leakage based on general fan performance curves and typical duct systems. The implications of the current European Community recommendations regarding nitrous fumes are also briefly considered. The actual problem relates to a 2100 m long decline that may have to be extended to 3000 m long without a connection to surface or through ventilation and is located in a hot tropical area where the surface wet bulb temperatures may exceed 25¦C for more than a third of the summer period. Once the preliminary ventilation design is completed, the auxiliary ventilation system can be modelled and simulations undertaken that incorporate heat and moisture transfer routines with the objective of determining the cooling load and amount of refrigeration required. Cognisance is taken of cyclic diesel operations and the thermal flywheel effect of the surrounding rock on the wet and dry bulb temperatures resulting from these operations.A suitable modular cooling system located at the portal is described and optimised with capital and operating costs of both the ventilation and cooling systems provided.

Jan 1, 2005

By Jed H. Mosgrove, Paul M. Budzak

INTRODUCTION Good mine ventilation starts at the drawing board. A coal mine will either be adequately ventilated or not, depending upon how good a job was done in the planning. Poor planning will not only result in inadequate ventilation, but may also lead to low productivity. A mine foreman cannot achieve high productivity while trying to maintain 4.2 m3/s (9000 cfm) of air at his last open crosscut. PLANNING What will provide adequate ventilation? A rule of thumb in laying out a coal mine is to provide enough entries in the main entry system to keep air velocities at 2.5 m/s (500 fpm) or less; at this rate of flow, good ventilation may be expected. The intake and return air courses of a coal mine do not improve with age. Entries will deteriorate, ribs will slough, roof will fall, and clutter will accumulate. The resistance offered to air travel will continue to increase with each passing year. Providing enough airways for the present and the future cannot be emphasized too strongly. In order to better appreciate good planning and bring it into better focus, assume a hypothetical mine has the following features: entries 1.5 m (5 ft) high and 6.1 m (20 ft) wide; one intake and one return entry; the mine development is 4.8 km (3 miles) from the intake and upcast ventilation shafts; and the ventilation requirement is 141.6 m3/s (300,000 cfm). Air travel of 4.8 km (3 miles) into a mine and the same distance back to the fan is not unusual. However, to do so with one intake and one

Jan 1, 1981

By Jed H. Mosgrove, Paul M. Budzak

INTRODUCTION Good mine ventilation starts at the drawing board. A coal mine will either be adequately ventilated or not, depending upon how good a job was done in the planning. Poor planning will not only result in inadequate ventilation, but may also lead to low productivity. A mine foreman cannot achieve high productivity while trying to maintain 4.2 m3/s (9000 cfm) of air at his last open crosscut. PLANNING What will provide adequate ventilation? A rule of thumb in laying out a coal mine is to provide enough entries in the main entry system to keep air velocities at 2.5 m/s (500 fpm) or less; at this rate of flow, good ventilation may be expected. The intake and return air courses of a coal mine do not improve with age. Entries will deteriorate, ribs will slough, roof will fall, and clutter will accumulate. The resistance offered to air travel will continue to increase with each passing year. Providing enough airways for the present and the future cannot be emphasized too strongly. In order to better appreciate good planning and bring it into better focus, assume a hypothetical mine has the following features: entries 1.5 m (5 ft) high and 6.1 m (20 ft) wide; one intake and one return entry; the mine development is 4.8 km (3 miles) from the intake and upcast ventilation shafts; and the ventilation requirement is 141.6 m3/s (300,000 cfm) . Air travel of 4.8 km (3 miles) into a mine and the same distance back to the fan is not unusual. However, to do so with one intake and one

Jan 1, 1981

By G. Reuben Yourt

IN 1935, Wright-Hargreaves Mines, Limited, initiated an enlarged programme of ventilation. In the mine, surveys were made from time to time by the geological staff, and improvements in ventilation were effected to accord with progress in mining as greater depth was attained. In the summer of 1938 an engineer was trained in the technique of dust sampling. Instruments and equipment for this purpose, such as konimeter, microscope, furnace, etc., were purchased, and a suitable office, equipped with laboratory facilities, was provided. The full time of the engineer is devoted to ventilation and dust control. On surface, in the mill crushing-plant and in the assay office and sample crushing room, an elaborate arrangement of hoods and exhaust systems has been installed to eliminate the dust. These installations, however, are not described in the present paper, which deals only with an extensive programme of underground ventilation as a method of coping with high dust concentrations, increasing temperatures and high humilities.

Jan 1, 1940

By Clifford S. Gibson

THESE troubled times suggest three reasons for maximum effort to control silicosis: (1) To keep to a minimum the ever-increasing burden occasioned by past deficiencies in this respect. (2) To effect the greatest production with maximum efficiency, which means at the least cost. (3) To preserve man-power in every way. Good control of the silicosis problem aids these aims. Inadequate control hampers all three. Insofar as our knowledge goes, silicosis in Ontario mines is caused by the inhalation of finely particulate quartz. It follows, therefore, that removal of dust containing this mineral automatically erases the cause. The achievement of this ideal is the aim of the mine operator. His best tool to this end is ventilation. This paper discusses mine ventilation from the viewpoint of dust control and it is hoped that in so doing there may result a better understanding of the part played by air in silicosis prevention. No attempt will be made to treat the subject fully. Instead, there will be presented certain salient aspects culled from field observations, the data in all cases referring to gold mines, unless mention to the contrary is made.

Jan 1, 1942

By John Harmon

THIS paper was written-with two objects in view: (1) To describe in detail what has been done toward the ventilation of the main unit of the Portovelo mines and the results; (2) to give information that may be useful to anyone who may have to make a preliminary mine ventilation survey, with recommendations for the improvement of the existing system and finally the maintenance of the improved system.

Jan 1, 1931

By D J. Brake

The underground metalliferous mining industry continues to change under the influence of many factors including, to name a few, new mining techniques and equipment, new approaches to risk management, more stringent safety and health standards, new employment practices and orebodies become deeper and often lower grade. In addition, the same technological revolution and plummeting cost of technology that has brought the GPS, smart phones and digital cameras to the general population is also working its way through the mining industry impacting on communications, automation and process control as well as introducing major efficiency and productivity dividends. However, not all factors are positive including, in particular, the issues of carbon costs and electrical power. This paper explores some of the trends affecting ventilation in hard rock mines, some of the existing and likely future challenges and some of the potential future solutions to these issues.CITATION:Brake, D J, 2013. Ventilation challenges facing the metalliferous sector, in Proceedings The Australian Mine Ventilation Conference , pp 3-12 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 1, 2013

By C M. Stewart

Lithium battery electric powered vehicles and machines (BEVs) continue to gain traction inunderground mining use, with predictions they will dominate the underground light and heavy vehicle fleets for some mines within the next decade. There are clear advantages for ventilation systems through reduced heat output and the elimination of harmful exhaust gases and diesel particulates. However, the risks and effects of BEV fire emergencies for underground personnel and ventilationsystems are still largely in the early stages of analysis and understanding. This paper conducts an industry review of a selection of lithium battery fire incidents and research.A research paper review on lihthium fire behaviour, intensity and toxicity will be analysed and comparisons made with equivalent diesel machinery fires for context. Guidelines for mine firemodelling parameters will be offered to assist in predicting the behaviour, heat and emissions from BEV fires, and recommendations and conclusions will be drawn on the specific hazards of lithium battery fires compared to conventional diesel powered machines.

Oct 12, 2022

By A Hatt

This paper presents a case study of characterisation work undertaken at a greenfield exploration project to provide engineering and OH&S inputs for the design of a ventilation system. The paper discusses how comprehensive early characterisation of the ambient surface and underground environment including aspects related to heat, dusts, strata gas, and geologic/mineralogical factors enables incorporation of the right ventilation and non-ventilation controls at a study level to mitigate the risks identified.

Oct 12, 2022

By Robert W. Miller

There are many problems faced by ventilation engineers in deep underground mining operations, not the least of which is controlling miner exposure to radon gas and its daughter products. Radon gas is commonly found in uranium mining operations, but may also be present in other deep metal mines. For example, tin mines in England, iron ore mines in Sweden, gold mines in South Africa, and molybdenum mines in the U. S. have potential radon exposures. This is because uranium and accompanying radium ore are ubiquitous to the earth's crust albeit at low levels. The fact that the activity represented by one WL can be caused by a relatively low concentration of radon gas increases the difficulty of control. Since the source of the radon gas is usually widespread throughout a mine, local exhaust ventilation is not a viable control schema. The technique used to control exposure is then dilution ventilation and, in fact, huge amounts of air must be moved in order to reduce potential exposures to an acceptable level. An interesting comparison can be made of ventilation rates in different types of mines. It is estimated in modern coal mines, which are generally acknowledged to have high rates of ventilation, that about eleven tons of air are moved for each ton of ore mined. A typical operating uranium mine may have ventilation flows of 14-15 tons per ton of ore mined. This provides an idea of the scope and importance of ventilation in modern mining operations where radon is a hazard. Further pressure is put on ventilation engineers by the steady downward trend in exposure limits set by national and international standard setting agencies. Much of this tendency toward lowered standards is based upon longitudinal mortality studies of miner populations. Another important factor is the limited number of experienced miners available in the labor pool. For optimum production, it is important to have as many experienced miners underground in each shift as possible. However, the average daily exposure in a U. S. mine must be less than .3 WL to permit the miner to work underground for a full year. The ventilation system then must provide enough uncontaminated air to maintain the WL below the .3 TTL level to maximize production efficiency and minimize personnel turnover and the problems associated with it. Ultimately, the goal of the ventilation engineer and health physicist is to protect the working miner from harmful exposures based upon currently acceptable standards. U. S. Federal regulations require that in uranium mines all active work sites must be monitored every two weeks if they measure above .1 WL. Areas that have .3 WL ratios or higher must be monitored on a weekly basis until five consecutive weekly samples show the level has dropped below .3 WL. Also, exposure records must be kept for all individuals exposed to levels exceeding .3 WL. These requirements provide a strong economic incentive to have a ventilation system that minimizes exposure of any personnel. A good ventilation system requires careful planning, operation and backup in order to fulfill its mission of providing adequate clean air. Its proper operation also requires coordination with production personnel so it can be adapted as new areas in the mine open up and old areas are sealed off. The ultimate indicator of ventilation efficiency to control radon daughter exposure is, of course, monitoring working levels. Historically, this has been done using the Kusnetz, Tsivoglou, and Rolle's methods, among others. These methods all require cumbersome equipment and tedious calculations to obtain the measurements that results in WL. More important, however, they require a significant time lag between sampling and counting, typically 40-90 minutes. This time lag is, in fact, what can cause significant economic losses due to unnecessary downtime as well as high WL exposures. In a typical mining situation, a sampling technician using the Kusnetz method takes a sample, moves to the next location and takes another sample and so on. Forty to ninety minutes after the first sample, the technician will stop, run the activity count on the filter and calculate the WL. The technician may be one-half mile away or several levels removed from where the first sample was taken when it is counted. If the WL ratio is high the technician must then backtrack to the sample position. There are then two options. If the sample area is a working stage, it can be shut down or a second sample can be taken. If the first alternative is chosen; i.e., shutdown and correction of the ventilation, then another sample must be taken, followed by a forty minute wait for results. If the ventilation adjustment didn't correct the problem, then the whole process must be repeated with a minimum of forty-five minutes per sample cycle when using the Kusnetz method. It has been estimated from operating uranium mines that the cost per hour for downtime on a production slope is about $1,50O/hour. The time lag between sampling and resultant data can be very costly. If the second alternative is chosen to verify the first reading, the miners may be unnecessarily exposed to high levels while waiting for the result. Clearly, such a sampling system can be markedly improved by eliminating the excessive time lag between sampling and analysis.

Jan 1, 1981

"An innovative concept is proposed whereby the operating power for haul trucks in certain areas of an underground mine are reduced in order to lower ventilation costs. Airflow that must be provided for diesel-powered equipment in most jurisdictions are regulated based on engine peak power capability and engine emissions, both of which can be lowered by artificially reducing the capacity of the engine through the engine control unit. Advances in asset tracking and ventilation on demand enable this methodology and this study demonstrates that it could represent significant value for mine operators. An examination of haul truck duty cycle indicates that de-rating of engine operating power based on the truck’s location in the mine can enable a step change in ventilation rates, often with no material impact on working conditions. Methods for implementation and estimate of cost savings will be defined.INTRODUCTION Ventilation of underground workings remains a significant operating cost burden for modern mine operators. In particular, ventilation rates for underground diesel mining equipment are often defined by local regulatory bodies and are assumed to present minimal opportunity for optimization or savings, at least until battery propulsion systems can be adopted. Commonly, the quantity of air flow required is defined by the peak rated engine output power of diesel equipment. For example, Ontario Regulation 854 SS 183.1 (3) states that “The flow of air must be at least 0.06 cubic metres per second for each kilowatt of power of the diesel-powered equipment operating in the workplace”. Jurisdictions that define air flow using the prescribed CANMET or MSHA ratings use peak engine emissions to calculate air flow required for dilution ventilation. These approaches are similar in that the maximum power capability of a vehicle’s engine is used to define the ventilation rate in all areas where it travels, regardless of what power output is actually used in any of those areas. Examination of the duty cycle of a typical underground haul truck identifies a large variance in the power output of the vehicle depending on its location in the mine and the activities undertaken at these locations. For example, travel on a mining sub-level requires a fraction of the truck’s peak power due to the minimal grade and significantly reduced travel speed. So if we can reliably reduce the peak engine power capacity and engine emissions in areas where peak power is not required, we can logically cut the ventilation rates for those areas."

Jan 1, 2017

The tangible effects of leakage are presented graphically, and a technique for quantifying the leakage characteristics of a ventilation column using the æSolveÆ capabilities of a common spreadsheet is demonstrated. A case study of a working decline in Western Australia is used to demonstrate the potential for more realistic design of development ventilating systems, and for more cost effective management of the condition of the duct.

Jan 1, 2002