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By M. R. Bitarafan

Mining method selection is the first and most important problem in mine design. In this selection some of the parameters such as geological and geotechnical properties, economic parameters and geographical factors are involved. The mining engineer must balance the input parameters and select the method that appears to be the most suitable; mining method selection is both an art and a science. Distinguishing all parameters that have an effect in mining method selection and in the determination of weighting factors of any parameter is a difficult task in an ordinary environment. This paper attempts to determine important parameters that impact on the mining method selection and demonstrates the calculation of the weighting factors for each selected parameter. This study has been directed to the research of a methodology for the selection of an optimal mining method. Different fuzzy methods are presented as an innovative tool for criteria aggregation in mining decision problems. Two methods of fuzzy multiple criteria decision making are used for selecting the optimal method with which to mine anomaly No. 3 of the Gol-Gohar iron mine. This paper finally presents a comparison of the selection methodologies and further discusses the most appropriate method investigated.

Jan 1, 2004

By Clarence Hall, CHARLES F. MUNROE

Of the common causes of the larger mine accidents, such as falls of roof and coal, gas and dust explosions, mine fires, and the misuse of explosives, all of which are often closely related, each must be studied and fought in a manner peculiar to itself. The last mentioned, the misuse of black powder and other explosives, is sometimes considered the least important of these causes of mine accidents; but its impor- tance is much greater than the statistics indicate, for the reason that the misuse of explosives is the true cause of many of the fatal mine fires, gas or dust explosions, and falls of roof that are credited to other causes. Both the quantity of explosives used and the number of purposes to which they are applied are increasing. They are now made at about 150 plants, in different parts of the United States, and the product of a single year is estimated at nearly 500,000,000 pounds. Nothing in all this material is a safe or "safety" explosive when in the hands of a careless or ignorant person; and this is true whether considered in connection with the shipment or the use of these explosives. In addition to the large losses of life and property resulting from an improper use of explosives in mining, the recent statistics of the railway bureau for the safe transportation of explo- sives have shown more than 400 persons killed or injured and over $3,000,000 worth of property destroyed by explosives in transit by rail. The fact that three years of cooperative effort under the wise supervision of this bureau has reduced these losses to almost nothing encourages the hope that similar cooperative effort may likewise greatly lessen losses of life and property from the use of explosives in mining. The additions to the large death roll of our mines make a recurring appeal to the public for fair treatment of the coal-mining industry, and to the miner and the manager that they join in every possible effort for greater safety. It may never be possible under conditions such as exist to-day to prevent mine accidents entirely. Little can be accomplished in that direction by either the operators or the miners working alone, but experience in all countries shows that

Jan 1, 1911

By A. R. Chambers

Malagash received its name from the Indians many years ago on acco.unt of its turbid (milky) waters, although,' the red men did not appreciate the significance of this turbidity, and were, of course, entirely ignorant of the crystal beds of pure salt that lay beneath their feet. The salt lake of the Mississippian, or Lower Carboniferous period, in which these beds occur, was apparently a large one, for there is evidence of its existence at intervals all the way from Sydney, Cape Breton, to St. John, New Brunswick, 300 miles as the crow flies, and there is no knowing what its other dimensions may have been, as much of this horizon is now buried on the one side beneath the waters of the Gulf of St. Lawrence, and on the other underneath the Atlantic ocean. About midway of the area and underlying the counties of Colchester and Cumberland, there occurs a synclinal basin of which the Malagash salt outcrop forms only a small portion of its northern limit. This basin has been described very fully by A. C. Hayes. (1) The operations carried on at Malagash since Hayes wrote his report have brought to light very little in the way of new geological data, except to confirm his pre-dictions as to the probable faulting of the salt strata, and the effect which these faults would have on mining operations.

Jan 1, 1924

By L. F. Willmott

The Bureau of Nines at Morgantown, W. Va., has been investigating, on a pilot plant scale, various methods of gasifying bituminous coal with oxygen and superheated steam to produce synthesis gas, as part of the broader program of finding new uses for coal. Differences in the pressure gasification methods used have depended on differences in the manner in which the pulverized coal is fed to the gasifier. One of the methods of feeding the coal that was investigated was the continuous, integrated coal-water slurry system. Goal and water were first mixed to produce a slurry; then the slurry was pumped through a heated coil. to evaporate the water fraction; the resulting steam was superheated, and some of it was separated so as to adjust the ratio of the residual steam to the coal fraction and make the resulting mixture suitable for partial combustion with oxygen in a pressure gasifier. Both before and during this reported3 developmental work it was necessary to make an extensive literature search into the technology of aqueous slurries in order to establish design criteria for the pilot plant and to correlate results of various phases of the work with that of other investigators, The pilot work coincided with the commercial development of a long-distance pipe- line for the transportation of coal as a coal-water slurry, and this fact tended to broaden the scope of the pilot plant work and the extent of the literature search. This bibliography is the compilation of the material that was collected.

Jan 1, 1963

By B. J. du Plessis

Automation of the conventional grinding circuits of Palabora Mining Company (PMC) required that adequate instrumentation and a control system had to be put in place, after which extensive testing had to be done to evaluate different grinding control strategies. The objective was either to optimize the particle size from the grinding circuit, or to increase milling circuit throughput while maintaining particle size at existing levels. Prior to automation of the process as described below the circuits were controlled manually. No on-line information on cyclone feed densities, cyclone feed flow or cyclone overflow particle size was available. This type of control relied heavily on interpretation of the process parameters by the operator and on his ability to make informed decisions on what the set points should be. An integral component of the automation process was the installation of a SCADA package (SysCad SDX) for implementing the on-line control strategy. A self-correcting model for prediction of cyclone particle size was developed and implemented on the SDX. The model is calibrated using values from the particle size analysers while they are operational. If the particle size analysers choke or fail for any reason, the model continues to provide a good estimate of the particle size. This technique ensures that a continuous particle size indication is available compared to the intermittent reading from the instrument. The continuous signal allows particle size to be used as a process variable for control purposes. A strategy was developed and implemented which allows for the production of a controlled grind size from the conventional milling circuit to the flotation circuit. The grinding control strategy allows for the control of particle size from the individual cyclone overflows. This strategy allows the operator of the process to increase or decrease production rate by specifying a coarser or finer grind size required for flotation.

Jan 1, 2001

By L. Nell

Bateman, like other engineering contractors, carries out or participates in many studies of potential projects in the Minerals Industry. This paper aims to explain the contracting approach to feasibility studies as practiced by Bateman, particularly where this differs from a traditional mining company approach. A brief description is provided of the usual phases of a project and the types of estimate needed to proceed from one phase to the next. Possible organizational structures involving different forms of association between owner, contractor and consultants are then outlined. The bulk of the paper then demonstrates that the cost estimates prepared stem directly from the mining plan, metallurgical process design and engineering design philosophy. The steps to be followed in order to prepare a credible estimate which accurately reflects the design constraints are described. The importance and the methodology of carrying out value engineering and of preparing contingency estimates based on a thorough risk analysis exercise are then discussed. Financial modelling techniques are described briefly. The paper concludes with short summary case studies of two recent investigations of potential projects in Africa.

Jan 1, 1999

By H. Naghavi, M. Karimi

"In self-aerated flotation machines, the gas rate depends on operational variables (e.g. froth depth and impeller speed), pulp properties (e.g. solid content and viscosity), and reagent addition (e.g. type and concentration of frother). The gas rate has a strong correlation with the flotation performance by influencing the gas dispersion properties and froth retention time. A factorial experimental design was used to study how the gas dispersion properties, the froth retention time, and the flotation performance respond to changes in froth depth and impeller speed (as the most common operational variables). An in-depth understanding of the effects of impeller speed and froth depth on the gas dispersion properties, especially the bubble surface area flux and froth retention time, is necessary to improve operating strategies for self-aerated flotation machines. All experiments were carried out in a 50 m3 selfaerated flotation cell in an iron ore processing plant. The results showed that the froth depth affected the metallurgical performance mostly via changing the froth retention time. The impeller speed had two important impacts on the metallurgical performance via varying both the froth retention time and the bubble surface area flux in the froth and pulp zones, respectively. The interaction effects of the froth depth and impeller speed were also established. This allowed us to develop a strategy for operating selfaerated flotation machines based on varying the froth depth and impeller speed with regard to the cell duty. IntroductionA mechanical flotation machine may be divided into two distinct zones. namely the pulp zone and the froth zone (Goodall and O’Connor, 1989; Yianatos, Bergh, and Cortes, 1998; Rahman, Ata, and Jameson, 2015a, 2015b). The overall flotation recovery, Ro, including the true flotation (particles attached to the bubble lamellae) and entrainment (particles recovered in water held in the bubble plateau boundaries), is a function of the recoveries in the two zones. It is calculated as follows (Dobby, 1984):"

Jul 1, 2019

By Urie R. W

Fluidized bed roasting of copper concentrates from Mt. Morgan, Mt. Lyell, Mt. Isa, and Peko was performed in a 3 ft 6 in diameter, slurry fed, pilot plant roaster operated under autogenous conditions to produce calcines with desirable leaching properties. A 12 in. diameter, externally heated roaster, provided with a dry feed system, was also used.An account is given of important factors which influence roaster behaviour and performance, and the effects of operating variables on calcine composition and leaching properties are described. Due to copper ferrite formation, a maximum of approximately 50 per cent of the copper in the calcines may be in the form of CuO before the copper leach efficiency falls significantly below lOO' per cent. Recovery of gold by cyanidation of leach residues was consistently high and independent of roasting conditions.INTRODUCTIONIn traditional practice copper is produced by pyrometallurgical techniques when the ,copper minerals in the ore or concentrate are predominantly sulphides, and by hydrometallurgy when the minerals are oxidized. Copper smelting is most suited to large scale operations, and for many years small to medium size mines with sulphide orebodies have been interested ill; alternative processes.Since hydrometallurgical plants (ie. leaching and electrowinning plants) are essentially unitized, they may be scaled down in output without much loss in efficiency or increase in processing cost, and many proposals have been made for the application of this type of process following roasting of the sulphides to render the copper soluble. However, until the advent of fluidized roasting, it was not possible to control the roasting operations sufficiently precisely to produce calcines consistently of the required leaching characteristics.

Jan 1, 1961

By J. H. Zietsman

It has previously been observed that industrial high-titania slags display a compositional invariance, remaining close to stoichio-metric M3O5. The present study focuses on this invariance and the phenomena which may cause it. The invariance was studied using thermochemical calculations and a mechanism is proposed for the observed behaviour. The mechanism suggests that a combination of thermochemical equilibrium and transient slag solidification and melting results in the observed compositional invariance. The proposed mechanism provides new insights and prompts questions about other mechanisms in the smelting process.

Jan 1, 2004

By Kung-Lee Wang

This report describes the linkage of all mining sectors with all other manufacturing, services, and final demand sectors of the U.S. economy. Besides being a basic data source, it provides a framework for aggregate economic impact analysis with respect to the mining industries. The framework is an open-static input-output model specifying 48 mining industries as classified in the Standard Industrial Classification Manual for all mining activities and 76 nonmining industries which collectively exhaust, all economic activities. The 124-sector model used in this report is an expansion of the preliminary 82-industry aggregation estimates made by the Department of Commerce. In the expanded model, 48 ruining industries are substituted for the original 6. The methodology used to ascertain the industry-by-industry procurement cost estimates and product ion flows necessary to the construction and expansion of the basic 82-sector model included personal interviews, case studies of specific mining activity analysis, and published and unpublished data from the Bureau of Mines and the Census of Mineral Industries, The end products of this study are two tables arranged to show the direct purchases necessary for each industry to support the level of economic activity and the total purchases generated in each industry throughout the economy by each $1 of aggregate economic demand in 1958. The tables may he used for economic impact studies and market analysis by firms engaged in or closely associated with the mining industries. Some illustrations of their usages are presented.

Jan 1, 1967

By Brendan Demich, Linda L. Chasko, Brendan Macdonald, Timothy J. Orr, Jessie J. Mechling, Cassandra Hoebbel, Jason Navoyski, William Helfrich, Jessica Cohen, Paul E. Schmidt, Jennica L. Bellanca

The National Institute for Occupational Safety and Health (NIOSH) in partnership with the Mine Safety and Health Administration (MSHA) developed VR Mine Rescue Training (VR-MRT). It is a training platform that allows mine rescue team members to explore virtual emergency scenarios collaboratively. Team members use a head-mounted display to visualize and interact together with equipment, hazards, and dynamic ventilation. All player actions are recorded and displayed for debriefing. Researchers at NIOSH demonstrated the training platform at various mine rescue competitions, including the International Mine Rescue Competition, and collected feedback. Participants’ responses were overwhelmingly positive. This paper describes the development of the software platform and the results of the usability surveys and interviews.

Jun 25, 2023

By Lowell B. Moon

The Stange and Byrnes Heirs manganese mines, adjoining properties in Bland and Giles Counties, Va., have produced at least 20,000 long tons of concentrates containing 8,000 long tons of manganese. They were selected as the site for a strategic-mineral development project because of the existence of geologic possibilities for extensions of the known ore body. In plan, the long dimension of the ore body parallels the axis of a major anticline. Localization apparently is due to longitudinal and transverse fracturing incident to two-way folding. That the manganese oxides were deposited by downward-moving solutions is suggested by the fact that mineable concentrations are confined to a coarse, porous sandstone under-lain by impervious silt and clay. Bureau of Mines project work began on January 15, 1941, and was completed on April 23. The work sampling the faces of the pit, bulldozer trenching, and churn drilling. Pit sampling consisted of 132 channel samples usually 5 feet long and taken at 20-foot intervals. Seventeen trenches aggregating some 4,500 feet in length were dug by the bulldozer. Trenches averaged 12.6 feet in width and 5 feet in depth. A total of 1,978 feet of churn drilling was done in 33 holes. In preparation for churn drilling it was necessary to construct roads to 21 of the 33 sites. A total of 2,867 linear feet of road averaging 13 feet in width was built.

Jan 1, 1950

By C. Blawert

In previous studies of various HPDC magnesium alloys we found that the corrosion properties of the skin are inferior in comparison to the bulk, However, this is surprising as one would expect better corrosion properties for the finer microstructure of the skin, In a more detailed study using 2 and 6 mm thick sections from HPDC AZ91 step plates, the corrosion results of the original skin, the slightly removed skin and the bulk were compared, Indeed, the corrosion measurements including potentiodynamic polarization, long term polarization resistance measurements and electrochemical impedance spectroscopy (EIS) revealed a lower corrosion resistance of the original as-cast skin. An explanation for this corrosion behavior was found, on the one hand, in the microstructure concerning especially the amount and distribution of the 13-phase and, on the other hand, in the composition, including the distribution of the most important alloying element Al and of some impurities such as Fe.

Jan 1, 2005

By X. Gumede

A model is being developed to quantify the benefits for mining houses resulting from the reduction in greenhouse gas and energy consumption by the application of innovative blast designs in addition to direct financial benefits throughout the mine to mill value chain. The conversion from shock tube initiation systems to electronics and some changes to powder factor measured on an open pit mine had previously been used to show significant improvement on blast outputs. Benefits were realised on drilling, blasting, loading, hauling and beneficiation. Quantification of the relationship between powder factor and reduced greenhouse gasses is investigated in this study. The study has highlighted new data that must be recorded and hence some assumptions were made as to fuel consumption and the energy in crushing was obtained from the Bond work equation. The mine to mill projects in an open pit platinum mine and a quarry produced energy savings and reduced greenhouse gas emissions associated with the careful application of explosives and blast innovations. The less fines project showed an increased carbon footprint. The model provides a new insight into how project feasibility should be decided based on economic and environmental viability.

Jan 1, 2012

By C. R. Edwards

"The Cigar Lake project will develop the world’s second largest high grade uranium deposit. The Cigar Lake Phase 1 orebody contains approximately 226 million pounds U3O8 at an average grade of 20.7% U3O8. The mine, to be operated by Cameco Corporation, will use a unique water jet boring mining method to cut previously frozen ore. A distinctive ore handling design is required to handle the high grade ore slurry exiting the mining machines. The ore will be crushed and ground underground, with the ore slurry then pumped to surface. Loaded into purpose-designed containers on tractor-trailer units, the ore slurry will be trucked to Cogema Resources Incorporated’s McClean Lake mill. McClean Lake will leach all of the ore. Forty-three percent of the resulting uranium solution will remain at McClean Lake for processing into yellowcake. The other 57% of the uranium solution will be shipped to Cameco’s Rabbit Lake mill for processing into yellowcake there. Detailed design for the Cigar Lake project is now underway. Barring regulatory or market delays, production could start up as soon as early 2005.INTRODUCTIONThe Athabasca Basin in northern Saskatchewan hosts a number of large uranium mining and milling operations, including Cluff Lake, Key Lake, McArthur River, McClean Lake and Rabbit Lake. (Edwards 1988, Edwards 1992, Fielder 1998, Edwards and Oliver 2000, Edwards 2001) The latest project in the Athabasca Basin is the Cigar Lake project."

Jan 1, 2002

By A. B. Cook, C. J. Walker, G. B. Spencer

Although water will displace oil from a petroleum reservoir to a greater extent than gas will, there are some reservoirs in which gas rather than water should be used for pressure maintenance. This is indicated by the high-percentage oil recovery in the case history reports on the Pickton field1 in East Texas and the Raleigh field2 in Mississippi. In gas cycling, total oil recovery includes vaporized products from the immobile oil in addition to oil produced by displacement. If a large part of the immobile oil is vaporized, total oil recovery may be higher than that obtainable from pressure maintenance by water injection. However, the amount of immobile oil vaporized may range from almost 0 to 100 percent, and no simple and reliable method has been presented in the literature for calculating oil vaporization. Development of such a method was the purpose of this study. Several papers concerning the calculation of oil vaporization have been published,3 6 and each is based on the concept dealing with K values — equilibrium constants for the various components in reservoir oils. (K for a component in a system of vapor and liquid in equilibrium is the mole fraction in the vapor phase divided by the mole fraction in the liquid phase.) This appears to be the most logical approach. Yet, the problem is too great for a perfect solution, even with modem laboratories to analyze oils and high-speed digital computers to perform calculations; the number of components in reservoir oils is too large. Because of their complexity no two reservoir oils are exactly alike. Also, K values change not only with variations in pressure and temperature but also with composition of the reservoir oil. In addition, during gas cycling, the lighter hydrocarbons tend to vaporize first. Thus, the reservoir oil becomes more dense and less volatile as gas cycling continues. Furthermore, the greatest amount of oil vaporization occurs near the injection well. Therefore, a simplified method is required for calculating oil vaporization because a rigorous method is not practical. The present methods of calculating vaporization generally require knowledge of the reservoir oil composition in which the mole fractions of the lighter components through hexanes are given; the remaining oil is described as C, + (heptanes and others of greater molecular weights). Assigning one K value for the C,+ system provides the simplest solution, but this incorrectly assumes that K for C, + does not change as gas cycling continues. This assumption can cause large errors because the K value may become less than a thousandth of the original after a large amount of oil vaporization. Another way to solve the problem is to determine changing K values for C, + according to how much of the C, + has been vaporized. To determine these values a sample of the reservoir oil is injected into a pressure-volume cell, and dry gas is batched in and out of the cell. Appropriate measurements and tests are made to determine K

Jan 1, 1970

By G de Jong, W Sunyoto, K Sari, P Silalahi

"An EXTENDED ABSTRACT is available for download. A full-length paper was not prepared for this presentation. The Big Gossan mine is operated by PT Freeport Indonesia and is one of the largest stope-and-fill mining systems in the world. The Big Gossan deposit is a ~200 m wide, 1200 m long and 500 m high steeply north-east-dipping to subvertical skarn body at the Cretaceous-Tertiary Ekmai-Waripi contact. The copper-gold mineralisation, with a 1.3 per cent copper equivalent cut-off grade, occurs mainly proximal to the Ekmai-Waripi contact in the garnet-clinopyroxene prograde skarn at the base of the Tertiary Waripi formation and partly distal to the contact in the tremolite-actinolite retrograde skarn of the Tertiary Waripi formation.The Big Gossan stope mine is designed from 2540 m to 3020 m above sea level. Present active stoping operations occur at level 2540 mL to level 2640 mL and are done in a primary and secondary stoping sequence. The width and length of the stopes is 15 m and 20 m respectively, while the height of the stope varies between 20 m and 40 m, depending on the ground conditions. The stability of the stopes is visualised in the Mathew’s Stope Stability graphs. These graphs and actual Cavity Measurement Scanner measurements on the stope will predict and show the overbreak and underbreak, which affect the reconciliation of the mined stope and also the pillar stability between the primary stopes. Geological factors like major structures, joints, veins and alteration intensity are important factors that correlated with the stability of the Big Gossan stopes.CITATION:Sari, K, de Jong, G, Silalahi, P and Sunyoto, W, 2015. Stope stability at the Big Gossan skarn deposit, Papua, Indonesia, in Proceedings PACRIM 2015 Congress, pp 139–146 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Mar 18, 2015

By D. P. Roberts

The stoping of platinum orebodies in the Bushveld has in the past been accomplished by using in-stope pillars to ensure stability of the workings. In the future, the use of continuous mining machines to mine the platinum orebodies is probable. It is not envisaged that such machines will cut in-stope pillars. This paper describes the alternative support systems that will replace the pillars and the stope geometries that will be required for this type of mining. The paper describes how the support resistance provided by in-stope pillars in current mining was determined and then evaluates support systems that can substitute for these pillars. Conceptual layouts based on existing breast mining layouts are presented. The layouts consist of panels of up to 100 m face length between rigid strike stabilizing pillars. A number of design charts are then proposed that allow for the design of appropriate support systems.

Jan 1, 2005

This investigation was planned to determine:1. The precipitation-hardening properties of the copper-nickel-phosphorus alloys; and2. The influence of arsenic and bismuth, singly and together, on the mechanical properties of these alloys.It was carried out in conjunction with an investigation, previously published in these Prooeedings on "The Influence of Various Elements on the Mechanical Properties of Copper containing a small percentage of Bismuth," and also formed part of a study of the influence of bismuth, nickel and phosphorus on the properties of commercial arsenical-copper.The first part of the present investigation arose out of a discovery some years ago of precipitation-hardening in phosphorus-deoxidized coppernickel boiler tubes, containing approximately 2% nickel and 0·03% phosphorus, supplied to an Australian Railways Department. At that time there was little information available on precipitation-hardening in copperrich copper-nickel-phosphorus alloys, but at the end of 1935 a short paper on the subject was published by E. C. Rollason and T. G. Bamford(2) and at the beginning of 1940 a comprehensive account was published by D. K. Crampton, H. L. Burghoff and J. T. Stacy.(3) Although the investigation carried, out independently by the present author was not as comprehensive as that of Crampton, Burghoff and Stacy, the results are in agreement with those of the American as well as of the British authors; hence, only a summary of this aspect of the work is needed here to establish a basis for a study of the influence of bismuth and arsenic.PREPARATION OF TEST PIECESAll alloys were melted under charcoal in plumbago crucibles in an oil-fired furnace and were poured out of contact with air into oil-dressed moulds to give billets either 7· 5 in. long X 0.83 in. diam. X 1·3 lb. weight or 12 in. long X 1·125 in. diam. X 3·3 lb. weight depending on requirements (the larger size was used, for example, in establishing the annealing characteristics of the precipitation-hardening alloys). The raw materials were electrolytic copper of the composition described in the previous paper(I) 6% copper-arsenic, 14% copper-phosphorus, 1:7% copper-bismuth, and Mond nickel in the form of "shot." The method of producing test pieces from these billets and the annealing treatments ana tests to which the pieces were subjected have been described in the previous paper.

Jan 1, 1944

By Foster Fraas

The most important lithium resource in the United States is in the form of the feldspathoid mineral spodumene. This mineral has not been used to any important extent as a source of lithium salts or lithium metal, as other more easily decomposed minerals have b en available in sufficient quantity to supply the relatively small demand. The need for cheaper lithium salts for such promising new uses as air conditioning or dehumidification and of lithium feldspar for use in the glass and [ramic] trades has justified the recent work done by the Bureau of Mines on King available, through improved beneficiation and extractive processes, some of the lower grade but extensive spodumene deposits.

Jan 1, 1937