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By B. R. T. Simoneit

The conversion of organic matter to petroleum by hydrothermal activity is an easy process, occurring in nature in many types of environments. Geologically immature organic matter of marine sediments is being altered by this process in Guaymas Basin (Gulf of California), Escanaba Trough and Middle Valley (northeast Pacific), Bransfield Strait (Antarctica), and Atlantis 11 and Kebrit Deeps (Red Sea). Contemporary organic detritus and viable microorganisms are also converted in part to petroleum-like products by the same process when present to become entrained, as for example on the East Pacific Rise at 13o N and 21o N and on the mid-Atlantic Ridge at 26oN. The hydrocarbon products (methane to asphalt) generated in all these areas have been elucidated in terms of composition, organic matter sources, and analogy to reservoir petroleum. This petroleum represents a major input of carbon to the primary chemosynthetic productivity of hydrothermal vent systems and may be important to interactions with metals in hydrothermal ore formation.

Jan 1, 1991

By Allan E. Koski

Michigan is the second largest iron ore producing state in the United States. The Empire Mine, celebrating its 30th anniversary in 1994, is the largest iron ore mine in the state. The Empire Iron Mining Partnership is an iron ore mining, concentrating and pelletizing facility managed by The Cleveland-Cliffs Iron Co., a subsidiary of Cleveland-Cliffs Inc. The Empire Mine is located near Palmer, MI on the Marquette Iron Range in Michigan's Upper Peninsula. The partnership includes Inland Steel Co., LTV Steel Co., Wheeling-Pittsburgh Steel Corp. and Cleveland-Cliffs. Mine construction began in 1960. The first pellets were produced in late 1963. Successive expansions in 1967, 1974 and 1979 resulted in Empire reaching a present annual capacity of 8.1 Mt (8 million long tons) of pellets. In late 1993, a milestone was reached with the production of Empire's 152-Mt (150-millionth long ton) of iron ore pellets.

Jan 1, 1994

By H. Douglas Dahl

In order to provide a reference framework for discussing the applicability of drilling technology to shaft construction in coal mining ventures, it is meaningful to review the history and nature of United States coal mines relative to shafting. Coal mining in the United States is characterized by shallow depths relative to most of the coal mining world. While the trend is toward deeper and deeper mines, many new mines have shaft depth less than 800 feet, and a large number of operating mines are drift or outcrop mines utilizing shafts for ventilation and portals. In the past, shafts of 200-300 feet have been commonly constructed. The use of conventionally sunk shafts at present and in the past is due in part to the shallow shaft length required, favoring conventionally sunk shafts. Another factor is that large shafts are required for multipurpose and production shafts. A third has been the practice of utilizing shaft construction contractors who had been until recently very competitive with raise drilled shafts at shallow to moderate depths. Recently, however, the costs of conventionally sunk shafts have skyrocketed. Too, more shafts are being required. Several factors are responsible among which are: (1) The increasing depth of shaft required for new mines and for shafts in advance of old mines. This significantly affects unit costs, e.g. $/yd3,of conventionally sunk shafts. Moreover, more gas is encountered in the shaft sinking operation.

Jan 1, 1976

By I. H. Warren, F. A. Forward, D. E. Pickett, Ernest Peters, Alvin H. Ross, I. M. Masters, N. P. Finkelstein, R. Derry, A. W. Fletcher

In this section the chemical, physical, and mechanical principles of hydrometallurgical processes are developed and reviewed. Plant practice is not discussed other than to note the metals which are processed using the various systems that are available. These notations will permit the reader to refer to the appropriate section in Part II of the Handbook for specific details on a particular metal or mineral. The equipment used to accomplish the desired unit operation is de¬scribed. Comparative data are presented to illustrate performance characteristics of unit operations, process, and equipment. The economic efficiency of individual processes is outlined. Some of the chapters contain material that is covered in detail in other sections of the Handbook Such chapters are developed only sufficiently to indicate the specific application of the techniques to hydrometallurgical processes. In this section considerable cross-referencing to the other related sections has been done.

Jan 1, 1985

By D. A. Barr

The Galore Creek copper deposits lie in the Stikine River district of northwest British Columbia, a largely undeveloped region which possesses rugged mountain terrain with deeply incised valleys which support dense undergrowth. Timberline is normally at about 1300 m (4265 ft) and the higher levels contain glacial systems and large ice fields with peaks rising to 3500 m (I 1,483 ft). A location map of the Galore Creek area is given in Fig. 1. In the late 1920s, field parties under the direction of F.A. Kerr (1948) of the Geological Survey of Canada completed geological surveys of the district. Primary access for this work was provided by riverboats on the Stikine and Iskut Rivers. Kerr's party had to contend with extremely difficult travel conditions that required considerable backpacking, augmented by pack dogs in certain cases. Absence of ground control led Kerr to establish his own topographic maps. Among the geological data established was the recognition [ ] that small bodies of orthoclase porphyry in the district are often well mineralized with base metal sulfides. Copper mineralization in the upper Galore Creek valley, a tributary of Scud River which flows into the Stikine River some 80 km (50 miles) south of the small settlement of Telegraph Creek, was originally discovered during aerial reconnaissance by M. Morrison and W. Buckholz while prospecting for Hudson Bay Exploration and Development Co. Ltd. in 1955. Malachite- and azurite-stained out- crops in several parts of the Galore Creek watershed are clearly visible from distances of 500 m ( 1640 ft) or more. Claims were staked in 1956 and the copper showings were partly mapped, trenched, and investigated by shallow diamond drilling. Because of the remoteness of the area and higher priorities for expenditures elsewhere, no further work was undertaken and most of the claims were allowed to expire.

Jan 1, 1991

By AK Raina

Blasting is used in mining and civil excavations to break the rockmass for ore production or creation of space. One of the disadvantages of the method is that it causes breakage beyond the proposed excavation lines. Pre-split is used to control such breakage in rockmass. This paper introduces the use of digital image analysis technique (DIAT) to measure rock breakage and half cast factor (HCF%). A new factor, undamaged area (AUD%), is introduced to improve the prediction, owing to the shortcomings of the HCF%. A comparison of significant data of AUD% and HCF% showed that HCF% has a lacuna of overpredicting, whereas AUD% can be used effectively to address the issue. Artificial neural networks are used to solve the complexity of multiple variables involved in predictions and to ascertain the importance and sensitivity of such variables in prediction regimes. This method will help users to evolve blast-induced damage objectively. A comprehensive scheme of things for future developments is included, which can be helpful to researchers in defining their research in this area.

By Robert M. Stevens, Arnoldo Acuña

INTRODUCTION In the early 1950s, underground miners and tunnel¬ing contractors were experimenting with ways to supple¬ment or eliminate track-bound loading and hauling equipment. Diesel-powered crawler and rubber-tire mounted front-end loaders and trucks designed for use on the surface were being modified and tried with limited but encouraging success. Miners began working with interested manufacturers to explore the develop¬ment of vehicles specifically designed to meet the de¬mands and constraints of underground material han¬dling. By the mid 1960s, load-haul-dump (LHD) vehicles were firmly established as a fundamental part of what has become known as the "trackless" mining concept, used in many mines and many countries around the world. The flexibility, mobility, and versatility of these units have given the industry a useful tool and have added new dimensions to mine development and produc¬tion. Many old mines have been redesigned to accom¬modate these trackless vehicles, and few new mining plans have failed to find some use for the concept during some phase of development and production. By 1977, several manufacturers in several different countries were producing a variety of LHD vehicles to meet the growing worldwide demand. GENERAL DESCRIPTION LHD vehicles combine certain characteristics of conventional front-end loaders and dump trucks, spe¬cifically designed for materials handling in underground mining and tunneling. The design intent is to provide one vehicle with one man, with the vehicle loading itself, hauling the load over level or inclined haulageways, and dumping the load. A typical production cycle of the most popular units has the operator driving the vehicle forward, forcing the bucket into the muck at or near floor level, and using the tractive effort of the vehicle and the prying action of the bucket to roll out and tilt back the load as shown in Fig. 1. With the bucket rolled back into the carrying position on the boom and the boom resting on the main frame of the vehicle, the load is hauled (trammed) to the dump point as shown in Fig. 2. At the dump point, the boom is raised, the bucket is rolled forward, and the load slides out as shown in Fig. 3. After dumping, the bucket is rolled back to the carrying position, the boom is returned to rest on the main frame, and the unit is driven back to the loading point for another load. There are two other types of LHD units that are slightly different in the cycles that they perform. One uses a hopper arrangement carried on the front of the machine, with a small bucket mounted ahead of the hopper. Several passes with the small bucket load the hopper, which is dumped by opening its floor. Al¬though this concept is several years old, it has gained less than 0.5% of the total market. The most recent LHD concept differs in that it has one bucket nesting inside another. The smaller inner bucket is loaded first, and it is then tilted back and above the larger bucket. When the larger bucket has been loaded, it is rolled back, and the total load is trammed to the dumping point. The dumping operation tilts the buckets in the reverse order from that of the loading operation. The impact of this new concept on the mining industry has not been measured yet. This chapter concentrates on the LHD vehicles employing the cycle described by Figs. 1 through 3, since these LHD units constitute an estimated 99% of the units in current use. MAIN VEHICLE FEATURES In general, the LHD vehicles available from most manufacturers share the same basic features.

Jan 1, 1982

By Wayne J. Nordstrom, Dennis L. Murr, James E. Wennen

National Steel Pellet Company (NSPC) conducted pilot scale testing and computer simulation studies which indicates that the capacity of the secondary grinding mill could be increased and that concentrate grade improvement could be expected with improved classification in the grinding circuit. Based on these studies, a decision was made to modify one secondary ball mill line at NSPC to include both cyclone and screen classification with magnetic separation applied to the circulating load. This paper will briefly discuss the studies leading to the decision to modify the circuit, the design of the modified circuit and the initial start-up and operating experience.

Jan 1, 1997

By Jair Carlos Koppe, João Felipe C. L. Costa, Luis Eduardo de Souza

United Nations (UN) has been leading an international committee to define the guidelines in reporting resources and reserves. The system proposed by UN is very similar to the guidelines currently in use in Australia known as JORC guidelines (Joint Organisation Reserves Committee). The JORC guidelines show major differences in relation to the Brazilian system current in practice. The important aspects differing JORC from the current Brazilian guidelines for reporting reserves is the required sampling spacing. The JORC system requires the points of observation to be closer spaced than the Brazilian system for any class of resources. All classification systems share some common aspects in terms of defining the classes of resources based on distance separating samples and on the degree of confidence or accuracy associated with the results reported. Despite of being very clear in terms of stating sample distances, all the systems of classification do not provide clear definitions on how confidence limits should be calculated. This paper proposes a geostatistical framework to measure the uncertainty associated with the reported tonnages and grades. Results demonstrate that the error associated with an estimate can be approximated using multiple simulated deposits to define confidence limits. The methodology is illustrated in a large Brazilian coal deposit and proved to be appropriate to address the issue of quantifying the uncertainty necessary to be incorporated in the systems of reserves classification.

Jan 1, 2002

By D. Lee Petersen, Charles R. Nelson, Marvin Marshak

INTRODUCTION Construction is underway on a laboratory cavern 713 m below surface at the Soudan Mine, Tower-Soudan State Park, in northeastern Minnesota. The 13.9 in wide by 11.6 m high by 71.8 in long cavern will provide a laboratory space for basic research in physics, in which experiments can be conducted with greatly reduced interference from cosmic radiation. A first generation experiment, called the Soudan 1 detector, is currently operating in a converted drift 591 in below surface at the mine. The initial phase of the Soudan 2 detector will be an 1100-metric ton unit, which will occupy approximately a 19 m length of the cavern. Control and monitoring equipment will be in a prefabricated shop office enclosure erected in the cavern. The location and planned usage for the laboratory cavern present special excavation, lining, and outfitting requirements. Temperature and humidity control, a relatively clean dust-free environment, and a long service life are required. Low cost and risk to the owner were basic requirements due to the nature of the funding. The most common type of site selected to meet these two constraints is either an operational mine or a civil works facility such as a road tunnel. There are often serious problems for such laboratories because of ongoing operations such as blasting, hauling, etc. Most of these problems do not exist for the Soudan 2 laboratory because the mine is used as a historical site with regular tours only in the summer season. The laboratory owner, the University of Minnesota, and the park owner, Minnesota Department of Natural Resources, are both public agencies of the

Jan 1, 1985

By Kadri Dagdelen, Thys B. Johnson, Salih Ramazan

Long-term production scheduling design is a very important part of mining because it determines the economic outcome of a project There have been many attempts to solve the optimum pit scheduling problem, but there has been no success in developing a scheduling method to give optimum results in maximizing net present value of a mining project. Open pit production scheduling problem can be formulated as a mathematical programming problem and solved using Mixed Integer Programming technique. However, it is not possible to solve the formulation within a reasonable time since the number of integer variables required for the mathematical model is too great to do that by today's available computer technology. There- fore, a methodology is required to combine, or aggregate, the mining blocks and to decrease the number of integer variables in scheduling without losing the optimality. Therefore, the fundamental tree concept was developed in this paper to combine the blocks. A fundamental tree is defined as any combination of blocks such that: 1. the blocks can be profitably mined, 2. the blocks obey the slope constraints and, 3. there is no subset of the chosen blocks that meets 1 and 2. A Linear Programming (LP) formulation is developed as a mathematical model to find a set of fundamental trees that exist for a deposit. Since the blocks are combined to form the fundamental trees, the number of variables required for the scheduling model is decreased significantly. This decrease in the number of variables makes it possible to mathematically formulate the multi-time period scheduling problem as a Mixed Integer Programming (MIP) model. A case study is performed on a multi-mineral copper deposit in Peru, South America. NPVs of annual after-tax cash flows generated based on the schedule provided by the "Fundamental Tree Algorithm" is compared with NPVs resulted from the schedules coming from three of the computer packages presently available to the mining industry. These are Mintec's M821V, Earth- work's NPV Scheduler, and Whittle's Milava open pit mine schedulers.

Jan 1, 2002

By Charles M. Brinckerhoff

Everyone -- especially the engineer -- is aware of the great technological and social changes taking place in our world today, We realize that changes have been taking place since man's beginnings on earth, but the element we now face is the speed of change, rather than change itself. Developments in medicine, in transportation, in space exploration, in global communications, and in many other fields, are constantly being made known. These developments are affecting the lives of each of us, although some may appear remote and unrelated to us as individuals. New developments, almost on a daily basis, affect our work as well as our social, religious and educational lives. The effects are immediate and permanent.

Jan 1, 1967

By S. Ersayin, I. Iwasaki

Laboratory- and pilot-scale tests were carried out to study the effects of a magnetically enhanced hydroseparator on a magnetic taconite sample. The test parameters included feed flow rate, percent solids and magnetic field strength. Laboratory tests showed that magnetic iron losses could be effectively prevented by applying a magnetic field strength of 0.003 T (30 gauss) to a hydroseparator. A relatively uniform magnetic field within a pilot-scale hydroseparator was obtained by installing a magnetic grid within the device. The test parameters for the pilot-scale study also included the effects of magnetizing/demagnetizing the feed and magnetizing current. Magnetic iron losses were substantially reduced, even when the hydroseparator was operated at very high upward velocities. Results also show that increasing the field strength decreased the losses, but magnetic enhancement was successful even at the lowest field strengths, and the results show that magnetic iron losses were lower when the feed was magnetized. Key words: Hydroseparator, Magnetic field, Magnetic grid, Taconite processing, Magnetizing, Demagnetizing.

Jan 1, 2003

By Richard C. Sale

INTRODUCTION The topic of the paper is to discuss the use of micro computers in the construction industry. I will first discuss the present situation of the PCs in our industry followed by a discussion how Guy. F. Atkinson has integrated the PCs into it's organization. I will close with some general comments that any organization should review in order to enhance the use of their PCs both at the job sites and home office locations. PRESENT SITUATION OF THE PCs IN OUR INDUSTRY Today, the possible number of combinations that can be obtained when bringing together the available models of PCs, with their peripherals and software, is astronomical. With this array of equipment that can be bought, plus the new and possibly better equipment that is being put on the market almost daily, it is indeed a very difficult task for an organization to know how to invest its money in computer equipment. At the job sites, the work is fast paced with little time available to fully learn what the PCs can do to help with the required paperwork. I am still waiting for the day when someone has to tell the superintendent at the job site that he cannot start pouring concrete or shoot the next round because no one can get the computer to work.

Jan 1, 1987

By W. Ter-Minassian

Besides the question of the stability of the roof and walls, one of the main problems encountered with tunnelling and shaft-sinking through the soft ground of alluvial origin forming the sub-grade of most large towns is settlement. Public transport and road tunnels in town centres are particularly affected, as they cannot generally be designed at any great depth for several reasons, such as the fact that ventilation shafts and access to the surface by means of stairs, escalators, lifts, ramps and so on must be fitted into the limited space available and are therefore costly, and because of the difficulties of driving below the water table. It is for this reason, coupled with the need for wide radius curves, that tunnels now so often pass close under buildings and monuments which must not be exposed to any risk of damage from subsidence during excavation. Grouting is the usual method used for consolidating soils, improving their strength and making them less compressible, but it is itself liable to cause heaving and cracks in the buildings above and has many other attendant dangers, including continually revised estimates which are the reason for the often inordinately high costs met with in urban tunnelling to-day. The mechanics of the problem have been studied and it has been found that a fraction of the total settlement is

Jan 1, 1997

By W. A. Crosby, A. Bauer

The recent increase in size of open-pit mining operations has resulted in major improvements in efficiency that are beneficial to the mining industry. Higher bench heights, larger diameter blast-holes, and more powerful explosives have played an important part in reducing mining costs. These measures have, however, resulted in increased energy concentrations in the blast area, which can result in severe backbreak problems for final pit walls. If backbreak is not controlled, a decrease in the overall pit slope angle will ultimately be necessary, and consequences, such as decreased recoverable ore reserves and increased waste-to-ore ratios, will result. Greater amounts of loose face rock will be produced, and planned safety berms will be less effective or nonexistent. Hazardous working conditions can also result. Remedial measures, such as scaling large areas and using wire mesh or other artificial support, are very expensive and difficult to implement. There must be a trade-off between the money saved by using larger blasts, and the time spent to maintain pit wall quality. The best approach is to control blasting effects so the inherent strength of pit walls is not destroyed. Presplit and trim blasting are two control blasting techniques used to accomplish this. Both techniques are designed to create a low explosive energy concentration per unit of wall area at the perimeter of the pit. Energy concentration of the main production blast must also be controlled to avoid damaging the final pit wall. A low-energy concentration at the final wall can be obtained by decoupling charges, decking charges, using less powerful explosives, decreasing blasthole diameters, and changing the burden and spacing. Rock properties-fragmentation and the final state of the pit wall-influence the success of a blast. The most important properties to consider are: in situ rock strength, and the nature, frequency, and orientation of structural features. Since these variables cannot be controlled, they must be evaluated by suitable field tests, and then a control blast can be designed using controllable variables, such as spacing, burden, and hole diameter, so the blast will be successful for appropriate rock conditions. One main thrust of this article is to show how control blasts can be successfully designed for larger hole sizes, up to 381-mm-diam. While efforts to produce a good control blast in shovel/truck pits have been ongoing for some time, recent use of a modified presplit technique in dragline operations has helped reduce water problems, improve safety, and increase dragline productivity in a number of surface coal mines. This article presents examples on designing and using presplit and trim blasting for large hole operations in both shovel/truck and dragline mines. Presplit Blasting As introduction to presplit blasting, Fig. 1 illustrates a typical presplit blast layout using 102 mm-diam presplit holes for 381¬mm-diam production holes. For this type of blast, presplit holes would normally be drilled first, ahead of main production holes. The choice can then be made between loading and firing the presplit line or infilling the main blast. In the latter case, the presplit line would be fired instantaneously 100-150 milliseconds before the main blast. As shown in the figure, the presplit line is formed ahead of the main

Jan 2, 1982

By Jennifer B. Leinart

With the signing of the North American Free Trade Agreement (NAFTA), Canada, the United States and Mexico entered into an international experiment in economics amid a cloud of controversy that still lingers. Central to the controversy are the disparities in wages that exist among workers in the signing countries. Wage surveys completed by Western Mine Engineering show the degree of disparity that exists among the mining industries of those three countries. Labor costs constitute 30% to 50% of operating costs. So these wage differences are a matter of interest to the mining industry. It is unclear how much of a factor they are in determining where companies search for and develop new mines. That depends on where minable deposits are expected to be found. However, a deposit that may not be minable in the United States or Canada may be minable in Mexico because of lower labor costs. Some adjustments are anticipated for the Mexican mining industry if and when its wages begin to approach comparability with US and Canadian wages. Table 1 combines the results of four recent surveys. It lists benchmark wages for mining job titles common to all the countries. About 220 mines are covered by the combined surveys. Canadian miners' wages, converted to US dollars, are typically 75% to 95% of US wages, while Mexican miners' wages are less than 10% of those in the United States. Table 1 also shows that wage disparities are not confined to international borders. In fact, the difference in average wages between the US coal industry and the US metal and industrial mineral industries is greater than between the Canadian and the US mining industries overall. Table 2 shows that managerial, technical and administrative salaries show similar, but less pronounced differences among the three countries. Salaried personnel at Canadian mines earn 80% to 90% as much as their US counterparts, while those in Mexico earn 20% to 30% as much. Here again, average US coal salaries exceed average salaries for the US metal and industrial mineral industries.

Jan 1, 1998

By E. Molleman

This paper reports on testwork exploring the dissolution/reprecipitation reaction: ZnS + 2HCl ? ZnCl2 + H2S, as part of a process for the production of pure ZnS. In spheroidal form, this would constitute a premium fonn of zinc refinery roaster feed enabling the production of pure ZnO. The research focused on the influence of proton activity on dissolution of zinc concentrates from Brunswick Mining & Smelting, Red Dog and Tara. The results and thermodynamic data suggest that the three concentrates contained significant amounts of the more reactive mineral wurtzite, although XRD analysis showed mainly the presence of the mineral sphalerite.

Jan 1, 1997

By T. Eisele, T. Weldum, G. Rajala, H. Walqu, S. Kawatra

The throughput and energy efficiency of comminution circuits is reduced when ore characteristics change, causing bottlenecks in some portions of the circuit while leaving excess capacity in others. The results of analysis of three years of daily grinding circuit operating data for a full-scale plant has shown that the location of bottle-necks in the circuit will change depending on the nature of the feed ore. The circuit tends to be limited by the primary mill capacity when the ore hardness increases, and limited by the secondary mill capacity with softer ores. Two methods are considered for shifting load in the circuit to relieve bottlenecks by taking advantage of excess capacity in other parts of the circuit: 1) Addition of supplemental comminution operations, such as high-pressure grinding rolls, to process specific size fractions that are not handled efficiently by the primary mill; and 2) Use of screens to shift grinding load upstream from the secondary mills back to the primary mills. In the case of the plant examined in this study, it was found that 58% of the time there was sufficient excess capacity in the primary mills that the throughput of the circuit could have been increased by using screens to shift the grinding load from the secondary mills to the primary mills.

Jan 1, 2004

By H. P. Knudsen

In most open pit gold mines ore waste selection is based on estimates made from blasthole samples. These estimates frequently are biased. While it is true that high grades and "outliers" can induce a bias they may not be the only reason for the persistent bias seen at many mines. This paper focuses on a source of the conditional bias. In this paper it is shown that the poor blasthole sampling can cause a conditional bias in the estimates.

Jan 1, 1991