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By T. H. Adair

"IntroductionCONTINUOUS casting of ferrous and non-ferrous metals is not a new innovation. Patents describing machines for continuous casting are on file as far back as the year 1842. In the middle of the last century, Sir Henry Bessemer conceived and built a continuous casting• machine for the production of mild steel sheet.The production of billets and slabs of both non-ferrous and ferrous metals is being carried out by the continuous method on a fairly large scale on this continent and abroad. At present, the volume of production of •copper and aluminum and their alloys in billet and slab form is much greater than the continuous casting of steel products. The new installation of a continuous casting unit at Atlas Steels, Limited, is designed primarily for the casting of alloy steel billets and slabs.As an introduction to the art of continuous .casting of metals, it is interesting to examine a few of the many devices which have been developed. As previously mentioned, Sir Henry Bessemer constructed a machine for the direct casting of mild steel sheet. His discoveries and ideas of continuously casting steel formed the foundation for the work of other pioneers in this field, all of whom have contributed to the improvement of this art through the ensuing years. The machines which will be selected for examination are those of Bessemer, Poland-Lindner, Williams, Ennor, and Junghans. The last unit to be discussed will be a modification of the Junghans machine recently installed at Atlas Steels."

Jan 1, 1954

By G. W. Poling, E. M. Coelho

"Sedimentary phosphate ores make up approximately 87% of the world's known reserves and are generally more difficult to beneficiate than igneous ores. In almost all cases, desliming prior to flotation represents the major source of losses. Laboratory studies of selective desliming of sedimentary ores are presented. Some fundamental studies with pure mineral systems are also discussed. Selective desliming, especially using selective flocculation, gave the best performance. Possible industrial applications of this selective flocculation-flotation procedure are discussed.IntroductionPhosphate ores are a major source of raw materials used in inorganic fertilizers (which supply nitrogen, phosphorus and potassium). Approximately 96% of the total world production of phosphate is utilized in the fertilizer industry. Reserves and production data are available in the literature. Today deposits with grades as low as 4% P2O5 are exploited in a few cases. However, in some northern African deposits, ore grades are still as high as 33 - 35% P2O5. Hout pointed out in 1982 that approximately 200 million tonnes of phosphate ore were upgraded by flotation only.Although there is a large number of phosphate minerals (over 300, only those in the apatite series are important ore minerals in phosphate ores. The most common member in this series is fluorapatite – Ca5(PO4)3F. However, in sedimentary phosphate rocks, many different substituents occur in the apatite lattice. Carbonate-fluorapatite (or francolite) is one of the most common names used to describe sedimentary apatites. It can be represented by the empirical formula:"

Jan 1, 1986

By Malcolm R. Gent

"The presence in Saskatchewan of sixteen kimberlite pipes of probable late Albian age has been reported. An abundance of aeromagnetic anomalies possibly attributable to kimberlite pipes have been identified and it is anticipated that industry has acquired reasonable proof of the presence of more than seventy k imberlite pipes in Saskatchewan. Historical reports of diamonds, the discovery of kimberlite pipes and these aeromagnetic anomalies indicate the Phanerozoic Basin and some parts of the exposed Shield in Saskatchewan have an excellent potential f or the occurrence of economically exploitable diamond deposits. The presence of some igneous minerals found in some salt solution breccias also indicates the probable occurrence of extensive diatreme activity during the Devonian.A Laramide Orogeny-induced crustal hinging or arching of the eastern margin of the Phanerozoic Basin and probably dextral wrenching-induced arching and faulting within the basin are proposed as major factors in the emplacement of Cretaceous or younger kimberlites. Zones of crustal weakness between subcratonic Archean blocks and internal faults are also thought to be significant controlling factors in the location of kimberlites in Saskatchewan.Mapped aeromagnetic anomalies interpreted as being possibly due to kimberlite pipes generally tend to be concentrated in, areas underlain by a basement of or with components of Archean age. Within Saskatchewan, these anomalies occur within the probable Trans-Hudson Orogen zone; the Rae-Hearn Craton margin; and the northeast flank of the Elkpoint Basin. Recognized areas of potential f or kimberlite emplacement are the Phanerozoic basin margin, the western limit of the Superior Province adjoining the Churchill Province, and a swath approximately 150m wide along the northeast-southwest trending Shaunavon linear and other areas of inner basin arching."

Jan 1, 1992

By E. M. De Souza, A. M. Coode

"Since potash was discovered substantial rock mechanics research has been done to develop the technology needed to cope with the unique conditions in Saskatchewan mines, to assure general mine safety and to assess the problems associated with mining potash at depth.The practical result of this research was the development of multiple entry systems incorporating yielding pillars for access ways and the chevron type layouts for production panels. These developments proved the viability of extracting potash from between 950 m and 1100 m depth using a high degree of mechanization. High extraction ratios could not, however, be achieved using the developed technology.The presence of a competent salt back is critical to mining the potash, and underground practice is governed by the presence of anomalous geological zones. These factors may be associated with the occurrence of brine inflows. One potash mine in Saskatchewan ceased its underground operations and was closed due to flooding of the entire mine. Two other potash mines in Saskatchewan had significant brine inflows. The costs involved in brine pumping, equipment relocation, bulkhead construction, grouting, and in auxiliary operations can run into the tens of millions of dollars.At Central Canada Potash, the use of stress relief and yielding pillars in the design of the mining layouts has resulted in over-all stable underground excavations and safe mining practices. There are two types of production room s, the chevron and the hybrid, which result in a planned block average extraction ratio of 35010. Mining activity has resulted in brine ingresses in two areas, and additional operating costs have been incurred over the years. These mining areas have demonstrated abnormally high closure rates, and rapid back deterioration. A program was established in those two areas to ensure the practice of safe mining procedures.The short-term perspectives inherent in the program included the elimination of all uncertainties associated with the possibility of an increase in brine inflow leading to flooding, the elimination of the abnormal ground behaviour and ground deterioration in the inflow areas, and the reduction of the long-term adverse effects of mining on regional mine stability. The program was also designed to provide means for developing alternative safe mine layouts which would improve the exploitation of potash resources. Details of the extensive work performed during the program are found elsewhere (E. De Souza, 1988).Grouting and salt backfilling are viable techniques to control brine inflows in underground excavations in potash mines. Detailed instrumentation of the backfilled test areas provides information on the long-term behaviour and effectiveness of the grout/backfill system, allowing for rational decisions on the complete abandonment of these mined out areas, thus resulting in considerable operational cost savings.This paper is presented in two parts: the first addressed the problems of diamond drilling , grouting and backfilling in Panel 209 and the second, addresses the problems of instrumentation and rock mechanics."

Jan 1, 1990

By W. B. Morrey

"In recent years, the major effort in detonator development has been concentrated on ""non-electric"" systems. Existing types all involve some form of explosive reaction to transfer ignition energy from outside the borehole to the pyrotechnic delay element contained within the detonator. These may be based on low-energy detonating cords having grained or plasticized explosive cores, shock tube energy transmission principals and explosive gas ignition systems. Each has properties of internal and external effects not present in electric conductors.The MAGNADET System introduces transformer coupling into electric blasting circuits. Safety features with respect to current leakage, stray currents of AC or DC source, electrostatic energy and radio-frequency radiation are comparable to ""non-electric"" systems, with advantages in simplicity, speed of hook-up and range of application.The transformer device used is a small ferrite ring or toroid which is an integral part of each specialized detonator. The detonator legwires form the secondary side of the transformer winding. The primary side is a single wire loop common to all units within the circuit. Power is supplied from a frequency-tuned portable exploder with unique circuit-sensing features.The MAGNADET System is expected to have particular application to shaft sinking, raising and tunnelling where the protective features as well as speed and efficiency are critical factors. Applications of the 'MAGNA' booster in quarries, open pits and longhole deck charges offer control over vibration problems and elimination of surface noise.IntroductionFor many years, electric detonators and electric blasting systems dominated surface and underground applications where controlled delay times were required. Control over the instant of detonation by blasting machines or power-line switches is the major safety feature.Electrical energy exists in many forms. It may originate from many sources. Electrical energy can be difficult to contain. It may enter or escape through a variety of conductive paths. It may fluctuate, concentrate or dissipate to pose a most elusive potential hazard."

Jan 1, 1982

By Kurt D. Mitchell

"This paper discusses general application for A C variable frequency drives in the mining industry. A review of induction motors characteristics, adjustable frequency controller fundamentals, and three commonly used controllers is presented. General application guidelines for the three different types of controllers; variable voltage input , current source input and pulse width modulated are then reviewed in detail, and an example of the economic benefit of this technology is discussed. IntroductionSoon after static DC variable speed drives were well established in industry, a need for static AC variable speed drives was seen. The first technical papers describing possible systems appeared in the mid-1960s, with commercially available units arriving in the early 1970s. Without exception, static AC drives are substantially more complex than their DC counterparts, but they have some very desirable properties. As compared to DC motors, AC motors are less costly, physically smaller, more reliable, require less maintenance , easily sealed, have good load regulation, are able to run at higher operating voltages, and have little danger of over-speed runaway . These properties are particularly desirable in explosion-proof, submersible or difficult accessibility applications.Variable speed AC drives are now a proven technology. available in packages from Sub-horsepower to hundreds of horsepower, from 115VAC to 4160 VAC. This paper covers the principles of operation of three common types of controllers, application criteria, motor consideration and finally an example of how an AC variable speed drive can pay for itself in terms of energy savings."

Jan 1, 1986

By Y. M. Shneerson, E. E. Zhmarin, E. M. Vigdorchik

One of the Gipronickel Institute activities is research and development of customized leaching technologies. The work is performed in several steps. Leaching kinetics are studied on a lab-scale according to a special procedure. Experimental data are processed using mathematical models of batch and continuous processes. Reliability of kinetic data and preliminary calculations is verified experimentally in laboratory reactors of various capacities. If necessary, certain parameters of the mathematical model are adjusted after pilot plant tests in continuous mode ( 4-reactor train, each with 25 L capacity). The specified kinetic characteristics along with the mathematical model are used for further development of industrial leaching technology.

Jan 1, 2005

By Y. M. Shneerson, I. V. Fomenko, M. A. Pleshkov, E. A. Ospanov, A. A. Shakhalov, S. S. Naboychenko

Low-grade copper concentrates (with copper content less than 20%) often contain impurities such as arsenic, zinc, lead, mercury, antimony and bismuth. The pyrometallurgical treatment of such concentrates is associated with several problems, which will ultimately affect the cost and the quality of copper cathodes as well as by-products. SRC Hydrometallurgy proposed a complex hydrometallurgical technology which allows the removal of impurities and the enrichment of low-grade copper concentrates to a quality which will be more suitable for pyrometallurgical processing. Laboratory batch tests and pilot plant results for the pressure leaching steps are shown in this paper. The key parameters of the leaching steps were determined and optimized. The quality of possible by-products were also determined. The developed flowsheet includes two pressure leaching steps: pressure oxidation (POX) and hydrothermal alteration (HTA). Both operations are carried out at 170°C with subsequent flotation of POX and HTA residues and recovery of copper, zinc and lead into separate concentrates for further processing. The development of the flowsheet was based on the Balkhash Zinc Plant (Balkhash, Kazakhstan), which had been non-operational since 2009. The developed flowsheet for processing of low-grade concentrates will utilize 80% of the existing equipment of the Balkhash Zinc Plant, thus leading to a significant reduction in required capital expenditure. INTRODUCTION As the age of copper mining operations advance, copper grades continue to diminish and impurity levels increase. Therefore, copper producers are forced to look for new innovative hydrometallurgical technologies which are able to handle low-grade raw materials and solve problems associated with impurities in an economically justifiable way. New technologies allow to extract valuable components selectively and to process low-grade concentrates which had previously been marked as non-technological and economically unviable. Hydrometallurgical technologies for processing of copper-containing materials can be divided into two main categories: chloride and sulfate (atmospheric, autoclave and bacterial) based technologies. Since the 1970s, many autoclave techniques for the processing of copper-containing materials have been developed and implemented (Naboichenko, 2002; Dreisinger, 2005; Habashi, 2009; Shneerson, 2011). The mineralogical composition of the ore (or concentrates), geographical location of the deposit, distance from industrial areas, availability of existing treatment facilities, availability of skilled labor, cost of electricity and cost of neutralization are some of the factors which are usually considered when the appropriate technology is selected (Halbe, 2002).

Jan 1, 2019

By P. K. Kaiser, S. Maloney, S. Yazici

"This paper focusses on the strength or capacity of cable bolts and introduces a new perspective for cable bolt design by demonstrating that mining-induced stress changes may affect cables in a detrimental manner. The concept underlying a novel cable bolt pull-out resistanceor bond capacity model is described and its practical significance is demonstrated by application to a simple two dimensional scenario of a fall-of-ground.A number of practical implications emerging from the application of this model are discussed to provide guidance to ground control personnel jar improving cable bolt support systems in mining applications. These include cable layout optimization, use of cable pre-reinforcements, assessment of embedment length requirements, need for plated cables (or straps), use of bird-caged cables (or buttons), and grout quality issues.IntroductionDuring the last decade, cable bolting has become a standard support technique in mining for handling wide spans, reinforcing are passes, preventing or reducing dilution of ore in marginally stable stopes, and for many other rock support need s.In its infancy, many deficiencies in cable bolt supports resulted from improper installation techniques and, in particular, grout quality control problems and the use of empty breather tubes (Goris, 1990). The development of special grout pumps and fail proof placement procedures (Oliver, 1992; Bourchier and O'Flaherty, 1992) have eliminated many of these early problems. Nevertheless, there are still situations where cables seem to lose their capacity during mining, and in these cases the cables do not provide the expected and needed ground support. In some instances, cables perform extremely well, whereas in others the same support layout and quality of support installation seem to be insufficient, allowing falls-of-ground or progressive cavings to occur."

Jan 1, 1992

By J. G. Henning, T. R. Yu, J. E. Croxall

"Test work was carried out to investigate the thermal response of loose and solid rock under varying environmental conditions. The primary interest was to develop a technique that could be used to increase the thermal variation between loose and solid rock for better detectability of loose rock underground with a thermal imaging infrared scanner.A variety of thermal enhancement methods were evaluated both in the laboratory and in the mine environment. Air temperature variation created by the exhaust from diesel driven, mobile mining machinery was studied in detail. When sufficient hot exhaust was supplied to the air, unsteady thermal conditions became more pronounced on the rock surf ace. This allowed a greater percentage of loose rock to be detected by a high-resolution infrared scanner within minutes.The degree of thermal enhancement of loose rock was found to be dependent directly on the thermal conductivity of rock, looseness, air temperature change, and inversely on the size of loose, airflow rate and distance downwind from the heat source. It was also found that an air temperature rise of at least r c, or rock temperature differential of1°C was necessary for minimal thermal enhancement required f or infrared detection."

Jan 1, 1990

By J. D. Zhu

A numerical model for predicting microporosity formation in aluminum castings has been developed, which describes the redistribution of hydrogen between solid and liquid phases, the transport of hydrogen in liquid by diffusion and Darcy flow in the mushy zone, One of the key features of the model is that it uses a two-stage approach for porosity prediction, In the first stage, the volume fraction of porosity is calculated based on the reduced pressure, whereas, in the second stage, at fractions solid greater than the liquid encapsulation point, the fraction porosity is calculated based on the volume of liquid trapped within the continuous solid network, which is estimated using a correlation based on the Niyama parameter, The porosity model uses thermal data calculated using the commercial finite element package ABAQUSTM. The model has been applied to examine the evolution of porosity in a series of experimental samples cast using unmodified A356 in which the initial hydrogen content was varied from 0,048 to 0.137 [cc/100g], A comparison between the model predictions and the experimental measurements indicates good agreement in terms of the variation in porosity with distance from the chill and the variation resulting from initial hydrogen content.

Jan 1, 2005

By W. S. (Steve) Vaughan

The Origins of Joint Ventures in Canada?s Mining Industry ? Among the most important financing structures used in our mineral industry today ? Virtually unheard of in the Canadian industry prior to the 1970s ? Borrowed from the U.S. oil and gas sector where it came into prominence in the late 1930s ? The Securities Acts of 1933 and 1934 in effect killed the sale of (equity) shares for financing oil and gas wells and start up mining companies

May 1, 2013

By L. C. Galloway, G. L. Tiley

"Mine hoist brakes must meet the following statutory performance criteria:- There should be two independent braking systems; at least one should act directly on the hoist drum.-If one brake fail s, the others should be able to retard and stop the hoist safely before the descending conveyance, carrying the rated load, reaches any obstruction.- When all brakes apply, decelerations should not be so violent that men travelling in the conveyance are injured.- The brakes shall not overheat and fade during an emergency stop.This paper discusses design principles and capacities of caliper and disc type brakes, when applied to double drum and friction hoists of equivalent capacity.IntroductionThis paper is about the brakes that are used on mine hoists; their functions, especially in emergencies, typical design, and how they are operated, controlled, tested, and adjusted.There are two basic types of mine hoists:1. Drum hoists, where the ropes are attached to, and coil on drums. The simplest system has only one drum. Most such hoists, however, have two drums, attached to a common shaft. Ropes are wound on each drum in opposite directions, so that when one conveyance is being raised, the other is lowered. The conveynaces may travel in vertical or inclined shafts. In the latter case, the hoist s are referred to as slope hoists.2. Friction hoist s, where the headropes are not attached to the wheel, but are in contact with it for an angle of lap of, typically, between 180 and 195 degree s. Forces are transmitted from the hoist wheel to the head ropes by friction.The lower ends of the conveyances are connected to balance ropes , approximately the same weight per foot as the headropes."

Jan 1, 1986

By R. F. Favreau, R. A. Morin, G. W. Kuzyk, P. J. Babulic, N. J. Tienkamp

"Atomic Energy of Canada Limited is constructing an Underground Research Laboratory (URL) as part of a comprehensive program to evaluate the concept of nuclear fuel waste disposal deep in crystalline rock formations. Careful blasting methods have been used to minimize damage to the excavation surfaces. Good wall quality is desirable in any excavation. In excavations required for nuclear waste disposal, the objective will be to minimize blast-induced fractures which may complicate the sealing requirements necessary to control subsequent movement of groundwater around a sealed disposal vault. The construction of the URL has provided an opportunity for the development of controlled blasting methods, especially for drilling accuracy and optimization of explosive loads in the perimeter and cushion holes. The work has been assisted by the use of blast simulations with the mathematical model Blaspa.This paper reviews the results of a recent project to develop a controlled method of full-face blasting, and compares the observed field results with the results of a blast simulator called Blaspa. Good agreement is found between the two, and the Blaspa results indicate quantitatively how the blasting may induce damage in the final excavation surface. In particular, the rock in the final wall may be stressed more severely by the cushion holes than by the perimeter holes. Bootleg of the rock between the perimeter and cushion rows occurs when the burst out velocity imparted to it by the explosive loads in the perimeter holes is inadequate. In practice, these findings indicate that quantitative rock stress and rock burst-out velocity criteria can be established to minimize wall damage and bootleg. Thus, blast simulations become an efficient way to design controlled blasting and to optimize quality of the excavation surf ace."

Jan 1, 1989

By N. Kacira

"Canada's largest known chromite deposit occurs in a layered intrusion in the Bird River area of Manitoba. However, numerically most of the chromite occurrences are in the Appalachians. Few, apparently minor, occurrences are also known in the Cordillera.In the Appalachians, chromite occurs as lenses and pods in dunite; they are of ""podiform"" type. The individual concentrations vary in size from a few pounds to tens of thousands of tons. Locally, chromite concentrations form discontinuous zones extending several hundred feet along strike and presumably down dip, therefore resembling boudinaged layers. Chromite layering and lineation are commonly concordant to those of the host rocks.The Appalachian chromite occurrences are thought to have been formed by fractional crystallization and perhaps also by liquid immiscibility. Their present shape and distribution are largely the result of deformation mostly prior to complete solidification of their host rocks. Therefore, the attitude of layering, foliation, lineation and other features related to deformation provide valuable guides in predicting the probable distribution of chromite, and hence are helpful in exploration.Chromites fulfilling the requirements for metallurgical, refractory and chemical grades occur in the Appalachians. Although the Appalachian deposits are similar in size and grade to some of the chromite deposits being mined elsewhere, they are unlikely to be economic, at present, in the higher wage Canadian environment. However, in case of shortages in chromite ore supply which may result, among other reasons, from increasing tendency to further processing in the producing countries, the Appalachian occurrences may become sources of ore. Such eventual shortages combined with relative abundance of electricity in Canada and recent innovations in oxide ore reduction technology may hasten the time when ferrochrome manufacture based on local chromite sources becomes economically viable."

Jan 1, 1982

By D. F. VanDine, H. George

"The possibility of spoil pile failure at the Kara Kara bauxite strip mine, Guyana, South America, is of major concern to the mine operators. The current mining technique requires the continuous operation of an 11 ms bucket capacity walking dragline on top of 18 metres of freshly spoiled uncompacted overburden material consisting of varying mixtures of kaolinitic clay and quartzitic sand. A field biased geotechnical study of undisturbed overburden and corresponding spoiled material has identified site-specific, non-linear correlations between vane shear strength, standard penetration resistance and spoil pile composition. Readily available geological exploration borehole logs has enabled a composite clay isopach map of the entire mining property to be drawn and critical areas to be zoned according to expected slope stability. Stability has been assessed using simple hand calculation techniques. Use of the resulting stability map provides the mine operators with a means of forecasting the need for appropriate preventative measures at varying stages of mine development, throughout the life of the mine.A similar investigation procedure could be employed at many presently operating and future strip mines.IntroductionThe Kara Kara bauxite stripe mine is located at Linden within the bauxite belt of Guyana in South America. The mining property covers a total area of 88 hectares of flat to gently sloping ground of which some forty per cent is being actively mined or being preparatorily stripped of overburden, prior to eventual mining (George and Gordon, 1979). The regional climate is tropic and humid. Temperatures vary from 16°C to 38°C and yearly rainfall averages 2600 mm, with at least 55 per cent of this precipitation occurring during the 'wet season' months of May to July and December to January."

Jan 1, 1983

By Amyn S. Dahya, David J. King

"The object of the paper is to discuss the developments in Carbon-In-Pulp (C.I.P.) technology for gold recovery. The CIP process is comprised of four major operations:(a) adsorption of precious metals onto activated carbon;(b) elution (stripping) of precious metals from the loaded carbon;(c) electrowinning of the precious metals from the concentrated eluate;(d) regeneration of the eluted carbon before recycle to the adsorption process.A detailed review of the technological developments in each of the above operations is presented together with a brief theoretical background of the underlying processes. The advantages and disadvantages of the CIP process are discussed in comparison with conventional zinc precipitation. IntroductionThe carbon-in-pulp (CIP) process involves the use of activated carbon for treatment of pulps containing dissolved precious metals, and provide s an alternative to the conventional gold recover y processes using filtration or counter-current decantation in conjunction with the Merrill-Crowe zinc precipitation process. The commercial application o f the CIP process in gold processing technology is a fairly recent development and it is of growing importance around the world.The fact that activated carbon ha s the property of adsorbing precious metals from solution was first published in 1847. In the following years , a great deal of work was carried out to develop an over-all recover y process using carbon, but all attempts were uneconomical because the loaded carbon had to be smelted to recover the gold.It was not until 1950 that Zadra, at the United States Bureau of Mines Research Centre in Reno, Nevada, developed a technique for stripping the loaded carbon using sodium sulphite-sodium hydroxide mixtures and then recovering the gold from the stripping solution by electrolysis. However, this process was not effective for stripping silver and further testwork was carried out on alternative stripping techniques. In 1952, Zadra published a paper describing the use of a hot sodium hydroxide- sodium cyanide stripping technique, which was effective in stripping both gold and silver. This paper included a description and flowsheet for the CIP process which is essentially the same as we know it today. This technique was used in the first major CIP plant, constructed and commissioned in 1973 by Hornestake Gold Mining Company. Since then there have been numerous CIP plants built in the U.S A., South Africa and Australia."

Jan 1, 1983

By I. I. Bethcherman

Introduction 1.-lmportance of Physical Nature of Condensates The metals of commercial interest which have reasonably high vapour pressures below 1,200° C. are mercury, cadmium, sodium, zinc, magnesium, lithium, and calcium. With the exception of mercury, these metals occupy high positions on the electrochemical series and form compounds with very high negative free-energies of formation. It is highly desirable that such metals should not be produced in a form which is finely divided, on account of their reactivity with atmospheric gases. Powders of magnesium, calcium, and even zinc (blue powder) are dangerous to handle and cannot be coalesced by simple melting techniques. Vaporization in vacuo forms a convenient method of handling and purifying the alkaline earth metals in particular and a previous paper by the authors dealt with some of the factors controlling the separation of metals by this technique (12). An essential part of the operation is the condensation of the vapour to produce a massive deposit which may be conveniently melted and handled by established metallurgical techniques. This investigation sought to determine the conditions affecting the nature of solid condensates.

Jan 1, 1951

By D. L. Mckay

"Texasgulf's 25-foot-diameter Kidd Creek No. 2 Shaft is situated next to the company's No .1 Shaft, so that both shafts can share common surface and underground facilities. Shaft construction, covering a span of five years from collaring to commissioning , started by freezing the 50 feet of grey varved clay covering the bedrock. After first completing the concrete headframe, sinking to the 5/00-foot horizon utilized two different shaft construction methods. The top 2800 feet was excavated by slashing a 6-foot-diameter borehole and hoisting the broken material up the 3OO0-foot-deep No.1 Shaft; the remaining 2300 feet was excavated by conventional benching methods. Prior to installing the 5100 tons of galvanized shaft steel on the completion of the sinking operation, the sinker was able to install all thirteen armoured electrical cables in one pass with his modified galloway stage. All loading and lip pockets were installed along with the steel guides and bun tons during the equipping operation.IntroductionOn November 21,1978, Texasgulf completed the No.2 Shaft at Kidd Creek Mine. This shaft is the largest shaft (volumewise) in North America and is unique in many respects . The purpose of this paper is to outline some of the unique features of this shaft as well as to record how this facility was constructed.Soon after sinking No.1 Shaft, Texasgulf started a program of exploratory diamond drilling from the 2800 level in No. 1 Mine. The drilling outlined sufficient ore below that level to justify a second mine. Feasibility studies demonstrated that production could be increased by one-third exclusively from this deep ore . The study called for a 25-foot-diameter shaft to the 5100-foot horizon to service a mine producing 5,000 tons of ore per working day.The site chosen for No.2 Shaft and head frame is 350 feet southwest of No.1 Shaft. This, along with No.1 Shaft, will be away from the influence of ground subsidence when mining at depth. The collar and head frame foundations are located on a flat rock ledge some 70 feet below surface. This flat ledge was found by test holing through the overburden by diamond drilling. It is fortunate that this location best satisfied the surface layout and logistics for both surface and underground ore and waste handling."

Jan 1, 1981

By K. Sun, S. A. Argyropoulos

The assimilation and the exothermic behavior of Zn-Zr metallic powder compacts into molten aluminum was studied by measuring simultaneously the temperatures inside at the centerline of the compact, at its surface, and inside the molten aluminum bath, as well as the net downward force acting on the compact. In addition, the exothermic behavior of these compacts was also investigated in an argon protected atmosphere. The temperature rise inside the compact increases significantly with the mass ratio of zirconium in the compact and to a lesser degree with the aluminum bath temperature. Experimental results on the recovery of these compacts in molten aluminum will also be presented.

Jan 1, 2008