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A subsurface study of the Maria Lode and associated hydrothermal alteration of andesites within the Karangahake epithermal gold-silver deposit defined the local evolutionary sequence in the once active geothermal system.   Some implications for mining of epithermal deposits arise as a result.

Jan 1, 1984

By C. L. Colburn

Discovery is the first step in the location of the claim. The provision of the Revised Statutes (2320) is, " No location of a mining claim shall be made until the discovery of the vein or lode within the limits of the claim located." Strictly speaking, discovery may be defined as the finding of ore or metalliferous rock in place in a well-defined vein. This definition was too strict, so discovery extended later to cover such indications of the pregence of ore, within rock in place, as would justify an experienced miner in spending his time and money with a reasonable expectation of finding ore in paying quantities. According to Colorado statutes a discovery must show a well-defined crevice; in Arizona, a lode, deposit, or mineral in place; in Montana, a well-defined crevice or valuable depositi; in Nevada, a lode or deposit of mineral in place. It is wrong to require a discovery. An honest prospector, wishing to obey the law, finds himself compelled, in most cases, to evade this point. In Clear Creek district, Colorado, a quartz seam will constitute a discovery; in Butte, a red capping; in Leadville or Ely, it is any place where a man can find a piece of vacant ground big enough upon which to sink a . 10-ft. hole. If the vein or deposit outcrops a man can show a discovery, but if it does not outcrop it is foolish to compel him to show one; and this point has been gently side-stepped. Any hole 10 ft. deep now constitutes a discovery. It is no use having laws that the people cannot consistently ob_ey. No sane man will dig unless he expects to find mineral, and he should be protected while digging. He should be required to show that he has mineral ground only when he comes to patent.

Jan 1, 1915

By J. A. Bullington

Mr. Chairman and Gentlemen: Some of you men may know who are familiar with what we have been doing in the C. F. & I. the last three years, that we have revamped our entire method of cost accounting from the pit bosses' report through to the final report of the president. The basic ideas involved in our cost system are not ours. They were given us by one of the best management and industrial engineering concerns in the country-at least they have that reputation. All I can do now is to try to pass on to you some of the ideas we have from various sources and the methods we have used in teaching those ideas to our coal men on cost accounting. This will not be a technical paper. I tried to put it in language we could all understand, and I hope I will be talking the same language that you gentlemen talk when you think about the cost of your coal in your own mines. Accounting of all kinds is classed as overhead expense, and as overhead expense we must justify its cost by the service it gives the rest of the organization. That brings us up to the problem of making the cost department pay its own way, just the same as the power plant has to pay its own way, and we have gotten away from the old idea of cost accounting in that it is sufficient within itself. The cost sheets are only so good as they enable the operating man to reduce his costs. The cost sheets are only just so much work for the cost department, and if they do not help the operating man to control his costs, we figure that is so much work wasted. The cost department works for the operating department.

Jan 1, 1931

By Franklin H. Allison

The influence of cobalt and nickel on the properties of steel might be readily expected to be very similar. The two elements occupy close and somewhat unusual positions in the periodic table, their chemical behavior and compounds are so completely analogous and their various physical properties, including the ferro-magnetic characteristic, show such 3 remarkable correspondence, that there is no doubt of a strong resemblance existing between these two metals. However, the variety of properties that are obtained from steel is due largely to the fact that the element iron possesses allotropic forms. Alloying elements that affect these allot ropic forms, or alter the nature of the allotropic changes, exert a profound influence on the resulting steel. In the case of nickel and cobalt, although these elements are very similar chemically and physically, it is found that they differ distinctly in their individual effects on the critical points of iron. Both nickel and cobalt form a series of homogeneous solid solutions with iron at temperatures above 1500" F. The difference in their behavior does not occur until changes begin to take place in the solid state below this temperature. An inspection of the respective thermal equilibrium diagrams at the lower temperatures makes the opposite effects of nickel and cobalt clearly evident. Effect of Nickel and Cobalt in Binary Alloys with Iron The binary diagram of iron-nickel has been investigated by a number of workers. The lower temperature portion in Fig. 1 is reconstructed from the data of D. and H, Hanson.1 The diagram of iron-cobalt, shown in Fig. 2, is taken from the work of Ruer and Kaneko.2 Masumoto3 has recently discovered a lower critical point in cobalt and this has been added to the diagram. In both diagrams the space lattice structures as determined by Andrews4 have been included. It is interesting to note

Jan 1, 1927

By Yuuichi Umeki, Yuusaku Tominaga

In order to keep good air condition in any specific branch such as coal face and/or heading in an underground coal mine, effective factors of all the branches with respect to the specified branch have been investigated, using computer simulation. Main results of the paper are as follows; 1) A branch in a network is classified to one of three classes of effectiveness to increase, to decrease and not to change airflow rate in the specific branch by altering aerodynamic resistance in the branch. 2) Distribution of position of branches, belonging to each class, related to a specific branch is one of the macroscopic characteristic of a network which is the inherent property of every network. 3) A relation of the aerodynamic resistances of any two branches to keep a constant airflow rate in the specified branch is shown as a monotonous curve. 4) Ventilation transition diagram with the rate of airflow in the two branches as coordinates, and for different rates of airflow in the specific branch as the parameter, is found to be straight lines. 5) Existing range of airflow rates In the specified branch is given as a parameter in the enclosed area in the ventilation transition diagram. Moreover, an expert system to assist ventilation control using a transition diagram of airflow rate is introduced in this paper.

Jan 1, 1991

By Hormando Leocadio Júnior

A otimização operacional no setor de desbaste da Linha de Tiras a Quente da Usiminas, Ipatinga, proporcionou redução do tempo de processamento dessa etapa, criando-se condições para aumento das temperaturas dos esboços na entrada do Trem Acabador (TA). Desse modo, para manutenção das temperaturas de acabamento dos laminados dentro da faixa de controle desejada, seriam necessárias menores velocidades de laminação no TA. Com a finalidade de garantir temperaturas desejadas na entrada do TA, que possibilitem adequadas velocidades, foi desenvolvido um sistema de controle de temperatura de esboço, que é acionado e desligado automaticamente na região inicial (topo), na extensão necessária do esboço. Esse sistema permitirá uma ampla exploração da nova condição de desbaste em menor tempo, com menor variabilidade da temperatura de entrada no TA, prevenindo ocorrências de estiramentos e de carepa, por temperatura alta nos topos e baixas nas bases dos laminados, maior uniformidade de propriedades mecânicas dos produtos, assim como aumento na capacidade produtiva.

Jul 30, 2018

By Kenneth Trelewicz

This report describes the total effort that went into developing environmental test criteria for general mine instrumentation. The project was a two-phase effort. Phase I consisted of research and mine visits to develop an environmental profile for mine instrumentation. The profile was translated into general requirements and then into test requirements. A review of existing government and commercial specifications was performed to determine if they matched mine instrumentation test requirements. Finally, a set of environmental test procedures was developed for the testing of mine instrumentation. Phase II consisted of selecting and purchasing sample mine instrumentation and subjecting them to the previously developed tests. The purpose was twofold: (1) to evaluate the criteria developed for use as an instrumentation test document and (2) to determine the effects of the various criteria on a representative sample of both proven and newly developed instrumentation. The test criteria were then modified, based on the results of the testing, and are listed therein.

Jan 1, 1981

By HOWARD I. YOUNG

F. W. SPERR, Houghton, Mich.-Why does 10 to 20 per cent. of the ore formation have to he lost? It is not clear to me why it should not all be taken out, from the conditions as I saw them underground this afternoon. H. I. YOUNG.-The grade of our ore is so low that we cannot stand the expense of timbering to support the roof, which, in a sheet-ground mine, would be from 25 to 50 per cent. of our total mining cost at present. After the orebody is mined out, we can often recover 10 to 20 per cent. of the material that is left. F. W. SPERR.-I think it should be a very simple mining engineering problem to mine all of that mineral without the use of a stick of timber. Similar work has been done in the mines of Grand Rapids Plaster Co. The usual operation in deposits of this kind is to open up and excavate as far as the roof will stand; when it caves, start in and dig somewhere else. But in the case I referred to they learned that by working the ground more systematically and leaving in a sufficient amount of pillar, they could later draw those pillars and get them all, and avoid uncon¬trolled breakages which would be likely to occur if they tried to hold up the overburden permanently. From the point of view of safety and also from the financial point of view, I contend that it is better to keep the caving under control and make it do what you want it to do, rather than to take out what you can conveniently for the time being and after a while be compelled to let the overburden do what it pleases.

Jan 1, 1918

By Patrick A. Turner, Timothy J. Olson, Abraham J. Schutte, Roman van Ommen

"The use of classifying devices in grinding circuits dates back to the early 1900’s when classifiers were installed to close ball mills. The search for ways to further improve grinding circuits led to the use of hydrocyclones in place of mechanical classifiers in South Africa in the early 1950’s. Today Hydrocyclones are considered as the best classifying devices to close grinding circuits. After 50 years of development the hydrocyclone has changed significantly when compared to the designs of the early days, but the focus to circuit optimization must be expanded to an understanding of the integral system consisting of mill, cyclone feed pump and cyclone.The key to good cyclone circuit design and operation in primary grinding circuits will be discussed including individual cyclone design features such as cone angle or materials of construction, as well as orientation and manifold configuration.Cyclone performance issues as well as maintenance considerations will be detailed in an effort to achieve consistent cyclone operation with minimal coarse solids in the overflow. “The importance of constant cyclone feed conditions is unquestioned today. The challenge of Cyclone sizing is a large number of not fully known variables prior to startup. The most significant unknown variable is slurry viscosity. Because of this, engineers often specify a conservative cyclone system that produces a finer than required separation. However, this approach often leads to higher circulating loads and a higher overflow density in order to produce the required separation. The higher cyclone overflow density results in a less efficient cyclone separation. The average concentrator can pick up 2-4% finer overall grind by having a more aggressive cyclone operation.”"

Jan 1, 2003

By Otis Hutchins

IT is intended to discuss in this paper the protection appliances used for high-temperature pyrometer installations involving the use of platinum couples and describe some of the characteristics of a new carborundum protection tube. Of all the components that go to make a complete pyrometer installation, the one given the least attention and the one about which there is the least exact information is the outer protection tube. This condition is unfortunate as much, and in some cases even the success of the equipment as a whole, depends on the proper type of protection. In the majority of cases, the installation is allowed to operate without attention until some part of the apparatus breaks clown. The outer tube is the component that usually fails and, more often than not, the failure causes breakage of the porcelain protection and ruin of the couple. This condition should be recognized and studied as it has a very important bearing upon the upkeep cost of pyrometer equipment. The cost of the outer protection tube is small compared with the cost of the platinum couple and procelain protection and it would well repay users of this type of equipment to replace their outer protection tubes at certain definite periods, which experience shows is necessary to prevent destruction of the platinum. Speaking generally, platinum thermocouple equipment is used for measuring temperatures over 2000° F. (1094° C.), and is used extensively for the control of brick and pipe kilns, glass pot furnaces, glass tank furnaces, heat-treating furnaces, open-hearth furnaces, and forge furnaces. A satisfactory protection tube must be sufficiently refractory not to soften at the maximum temperature within the furnace. It must be resistant to cracking clue to temperature changes, should be as nearly gas-tight as possible, and should be made of a material with a high thermal conductivity and great resistance toward erosion by the furnace atmosphere.

Jan 9, 1919

The basic objectives of mine roadways are to provide sufficient cross sections in order to meet the needs of accommodation of equipment, transport, personnel travel and ventilation. However, many roadways become damaged to the extent of needing maintenance, generally dinting, and in some cases requiring re-ripping. The damaging of a roadway is a result of several factors including geological properties of surrounding strata, method of roadway formation, type and spacing of support and so on. Among these factors, strata conditions play an important effect on the stability of roadway. Weak rocks cause excessive roadway closures and water softens some rocks and promotes the closure problem. In this paper, a scientific discussion on the effect of water on the stability of roadways is given on the basis of results obtained by means of in-situ investigation and laboratory rock tests.

Jan 1, 1996

By Gökhan Çevikbilen, Ümit Karadoğan, Sevde Korkut, Berrak Teymur

This study proposes a methodology for dewatering of mine waste using a multistage process, combining geotextile tubes and anionic polyacrylamide (PAM), which accelerates the recycling process of water for aggregate mine waste facilities. For this purpose, waste from a rock quarry located in Istanbul was investigated. Flocculation tests revealed that the use of anionic PAM at an optimum dosage reduced sedimentation time and decreased turbidity. The compatibility of geotextiles and mine waste conditioned with different dosages of the effective PAM was evaluated by rapid dewatering tests, considering the water content of the filter cake, amount of infiltrated solid content, and rate of filtration. Geotextile dewatering tests were also performed on PAM-free and conditioned suspensions. The dewatering process was optimized by the use of a geotextile tube in two stages: (i) on the unconditioned suspension and (ii) on recovered PAM-conditioned leachate. Directing PAM-free non-hazardous soil-based material obtained in stage i to alternative beneficial uses is suggested, to minimize the volume of waste. Furthermore, recycling of the effluent obtained at the second stage of the proposed method will ensure more efficient water reuse

Sep 29, 2022

Work undertaken in the field of cut-off grade optimization has not advanced much beyond the work undertaken by Lane in 1988. His definitive work is based on the calculus of the Net Present Value criterion which is the most widely understood, consistent, and appropriate method by which sequential cash flows arising from the extraction of mineral reserves from an exhaustible resource can be represented. Although the mathematics is not complex Lane’s method is not a widely appreciated or applied approach to maximizing the value of a mining operation through selection and balancing of operational cut-off grades. Three stages in a mining operation namely, mining, processing and marketing were defined by Lane and the economics of each stage are identified and isolated to provide an optimum cut-off for each stage. Points of intersection along the present value curves for each stage of a mining operation are used to identify balancing cut-off grades at points where the capacity of the mining, processing and marketing stages is fully utilized and the Net Present Value of the operation is optimized. Data that simulate a small, deep-level Witwatersrand-type gold mine with an average grade of 6 g/t Au and a logrithmic variance of 1.2,are applied to the model to illustrate the benefits of representing cut-off grades in terms of present value curves for different segments of a mining operation. Keywords: Cut-off grade, Net Present Value, grade-tonnage curve, balancing cut-off grade, mining, processing, marketing.

Jan 1, 2003

By George Koenig

I have been granted the outstanding privilege... and pleasure to introduce you this evening to the Grand-Duchy of Luxembourg and to tell you in a few words how this country came into existence, how it managed to survive and what role it nowadays plays in the International Community. Luxembourg, situated between Belgium, Germany and Prance is a small, a very small country... just a dot on a map. After all, it is only a matter of 62 miles from top to bottom and 37 miles across at it's widest... barely covering some thousand square miles... so be careful about stepping on the accelerator... you might be in and out without having realized it... which would be a real shame as the Grand-Duchy enjoys a fairly unique geographical feature in so far as it provides the visitor with an outstanding variety of scenery concentrated in a minute spot.

Jan 1, 1987

By Alan R. Johnson

Because of the competitive nature of the iron ore pelletizing industry, commodities required for the production of pellets are continually being scrutinized to minimize their usage and offset rapidly escalating costs. In the Tilden pellet plant, the usage of one commodity, bentonite, has been reduced from over twenty-four pounds per long ton of pellets to as low as eight pounds per long ton. The evolution of bentonite usage from plant start-up to the present time is described.

Jan 1, 1984

By Stanley D. Michaelson

The selection of the proper size of grinding mill for a given installation has long been a subject of discussion by mill manufacturers, consulting engineers, and their clients. It would be presumptuous for the author to propose a simple cure-all. Instead this paper describes a method that has been used over a period of years to determine mill size with a reasonable degree of accuracy. Furthermore, this method claims no theoretical background, for ball milling, as practiced today, is still an art. The selection of the proper size and type of mill must, therefore, be made on an empirical basis. A ball mill, in order to grind, must deliver power through its grinding media to the material to be ground. Assuming that a mill is properly supplied with grinding media, ore pulp, and power, the amount of grinding done is closely proportional to the power expended. With this as a basic premise we can, for the moment, forget about type of mill (i.e., overflow, low-level mill, or other design) and base our calculation of mill size on the total power required to grind a ton of a particular ore from a given size of feed to a given size of product. When this is determined the type and size of mill, its speed, and the weight of the load of grinding media can be selected to absorb this amount of power. Development of Technique Some years ago, W. L. Maxson, F. Cadena, and F. C. Bond, working in the Minerals Research Laboratory of Allis-Chalmers Co., Milwaukee, developed and described a laboratory technique for comparing grindability (or relative ease of grinding) of various ores at different mesh sizes of grind.' Additional papers2 have described the work done since that time, and the most recent paper3 summarizes results for hundreds of tests on many ores and materials. The essentials of the testing method are: a 12 by 12-in. laboratory mill operating at a constant speed of rotation, using a constant volume of ore, and a constant ball load having a relatively constant grinding surface. Tests are run at 250 per cent circulating load with the feed, circulating load and products screened on calibrated Tyler standard screens. Certain ores from operating mills, representing various mineralogical types, were selected for reference and designated "standard ores." In most cases, detailed studies of the grinding circuits were made when samples of mill feed were taken for grindability tests. Each analysis formed a basis for comparison of the grindability in the laboratory with actual power requirements for a particular grinding circuit. In addition, a critical study was made to evaluate the differences in the various grinding circuits so that two or more field installations having different layouts (single-stage grinding, two-stage grinding, or others), and using mills of various diameters operating at various speeds with differing circulating loads, etc., could be compared. Large-scale tests by mill operators complemented by laboratory investiga-

Jan 1, 1947

By Stanley D. Michaelson

The selection of the proper size of grinding mill for a given installation has long been a subject of discussion by mill manufacturers, consulting engineers, and their clients. It would be presumptuous for the author to propose a simple cure-all. Instead this paper describes a method that has been used over a period of years to determine mill size with a reasonable degree of accuracy. Furthermore, this method claims no theoretical background, for ball milling, as practiced today, is still an art. The selection of the proper size and type of mill must, therefore, be made on an empirical basis. A ball mill, in order to grind, must deliver power through its grinding media to the material to be ground. Assuming that a mill is properly supplied with grinding media, ore pulp, and power, the amount of grinding done is closely proportional to the power expended. With this as a basic premise we can, for the moment, forget about type of mill (i.e., overflow, low-level mill, or other design) and base our calculation of mill size on the total power required to grind a ton of a particular ore from a given size of feed to a given size of product. When this is determined the type and size of mill, its speed, and the weight of the load of grinding media can be selected to absorb this amount of power. Development of Technique Some years ago, W. L. Maxson, F. Cadena, and F. C. Bond, working in the Minerals Research Laboratory of Allis-Chalmers Co., Milwaukee, developed and described a laboratory technique for comparing grindability (or relative ease of grinding) of various ores at different mesh sizes of grind.' Additional papers2 have described the work done since that time, and the most recent paper3 summarizes results for hundreds of tests on many ores and materials. The essentials of the testing method are: a 12 by 12-in. laboratory mill operating at a constant speed of rotation, using a constant volume of ore, and a constant ball load having a relatively constant grinding surface. Tests are run at 250 per cent circulating load with the feed, circulating load and products screened on calibrated Tyler standard screens. Certain ores from operating mills, representing various mineralogical types, were selected for reference and designated "standard ores." In most cases, detailed studies of the grinding circuits were made when samples of mill feed were taken for grindability tests. Each analysis formed a basis for comparison of the grindability in the laboratory with actual power requirements for a particular grinding circuit. In addition, a critical study was made to evaluate the differences in the various grinding circuits so that two or more field installations having different layouts (single-stage grinding, two-stage grinding, or others), and using mills of various diameters operating at various speeds with differing circulating loads, etc., could be compared. Large-scale tests by mill operators complemented by laboratory investiga-

Jan 1, 1947

By Stanley D. Michaelson

THE selection of the proper size of grinding mill for a given installation has long been a subject of discussion by mill manufacturers, consulting engineers, and their clients. It would be presumptuous for the author to propose a simple cure-all. Instead this paper describes a method that has been used over a period of years to determine mill size with a reasonable degree of accuracy. Furthermore, this method claims no theoretical background, for ball milling, as practiced today, is still an art. The selection of the proper size and type of mill must, therefore, be made on an empirical basis. A ball mill, in order to grind, must deliver power through its grinding media to the material to be ground. Assuming that a mill is properly supplied with grinding media, ore pulp, and power, the amount of grinding done is closely proportional to the power expended. With this as a basic premise we can, for the moment, forget about type of mill (i.e., overflow, low-level mill, or other design) and base our calculation of mill size on the total power required to grind a ton of a particular ore from a given size of feed to a given size of product. When this is determined the type and size of mill, its speed, and the weight of the load of grinding media can be selected to absorb this amount of power. DEVELOPMENT OF TECHNIQUE Some years ago, W. L. Maxson, F. Cadena, and F. C. Bond, working in the Minerals Research Laboratory of Allis-Chalmers Co., Milwaukee, developed and described a laboratory technique for comparing grindability (or relative ease of grinding) of various ores at different mesh sizes of grind.1 Additional papers2 have described the work done since that time, and the most recent paper3 summarizes results for hundreds of tests on many ores and materials. The essentials of the testing method are: a I a by 12-in. laboratory mill operating at a constant speed of rotation, using a constant volume of ore, and a constant ball load having a relatively constant grinding surface. Tests are run at 250 per cent circulating load with the feed, circulating load and products screened on calibrated Tyler standard screens. Certain ores from operating mills, representing various mineralogical types, were selected for reference and designated "standard ores." In most cases, detailed studies of the grinding circuits were made when samples of mill feed were taken for grindability tests. Each analysis formed a basis for comparison of the grindability in the laboratory with actual power requirements for a particular grinding circuit. In addition, a critical study was made, to evaluate the differences in the various grinding circuits so that two or more field installations having different layouts (single-stage grinding, two-stage grinding, or others), and using mills of various diameters operating at various speeds with differing circulating loads, etc., could be compared. Large-scale tests by mill operators complemented by laboratory investiga-

Jan 1, 1945

WITHOUT in any way detracting from the credit due those engineer-miners of copper who operate with power shovels, it may be said that compared with block-caving underground their work is simplicity itself. True enough, ingenuity and foresight of high order are required to lay out an open-cut operation to the best advantage; good organization and intelligent care and use of machinery are required to achieve minimum costs. But really successful undercut block-caving demands all this; and, in addition, it requires a conception and application of unseen natural forces that reflect a type of genius that is rare. Assuming that preparatory work has gained access to the orebody, mining of whatever kind consists essentially of three operations: 1. Breaking the ore "in place," as Nature left it, into pieces small enough for convenient handling. 2. Loading these pieces into some kind of a vehicle on wheels. 3. Transporting the broken ore to some central plant on the surface where the process of extracting the valuable metals can be commenced. In shovel mining, dynamite and steam or electric power are used-with remarkable efficiency, be it said-to accomplish each of these three processes. In underground-caving methods the old force of gravity is called upon to accomplish, in large part, processes 1 and 2. To do this with success is an artistic achievement of a sort. Put it another way. The earlier methods of underground mining constituted a translation of surface procedure to an underground theater. The ore was blasted and dug out in small quantities; in its place were put props, usually of timber, to prevent the caving of the rock above; and men toiled with

Jan 1, 1933

By C. L. de Jongh

The number and scale of backfilling operations in the South African gold mining industry have grown rapidly in recent years. The application of backfilling as a means of stope support in a wide range of geological conditions and mining systems has resulted in the development of a variety of alternative fill systems. Experience gained in the evaluation of operating fill systems has contributed to a knowledge of the essential concepts to be considered in their design. This knowledge can also influence management's perception of what is expected of a backfilling system. The factors to be assessed in stating and satisfying these objectives are examined. The areas of investigation cover the selection of a suitable material, its preparation, distribution and placement, and the engineering logistics involved in the management of such a system. Finally, the integration of a filling system with an existing mining operation is examined.

Jan 1, 1988