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By F. H. Mertens

The costs of an open-pit operation can be classified according to the following hierarchy : -overall cost of production (per ton of ore) -costs of basic operations and functions -costs of auxiliary operations -costs of machines and equipment -maintenance costs. Production cost is a measure of the open-pit mine performance and should be the final target of the management control. It is also a measure of the share of the mine in the processing stages.

Jan 1, 1990

By Jose Maria Rodriguez-Ibabe, Ronaldo Antonio Neves Marques Barbosa, Douglas Glenn Stalheim

Major costs in the production of both flat and long commodity grade structural steel products include alloy, labor and energy. Flat and long commodity grade structural steels such as ASTM A36, ASTM A527Gr50, S235, S275, S355 and other equivalent world societal standards represent over 500 million annual tons worldwide. These simple commodity grades consist of a minimum of three base common alloying elements, carbon, manganese, silicon and then can be supplemented with microalloying elements of either vanadium or niobium. Since 2016 raw material costs for two of the five alloying elements in these commodity grade structural steels, FeMn and FeV, have risen significantly and/or have become volatile making it difficult to maintain stability in profitability for these commodity grades. For steel plants producing hundreds of thousands and in some cases over a million tons annually of these common structural steel grades because of the significant alloy cost increase for Mn and V alloy additions has squeezed profitability of these grades. These grades typically represent the base loading for cost controls in most all steel plants and hence a significant cost increase or volatility in two of the five elements used for these grades having a negative effect on overall production costs. However, with a proper strategy for alloy designs working in conjunction with the mills existing processing capabilities to achieve the desired end metallurgy/mechanical properties, alloy costs and operational efficiencies can be realized.

Oct 1, 2019

By T. W. West

A description of a coal washing spiral is presented, together with a discussion of the parameters which define the envelope within which the spiral can be successfully applied. The correct process engineering, installation, operating and maintenance parameters are addressed as well as actual results for new and retrofit installations on USA coal capital and operating costs of spiral circuits are compared to those of traditional fine coal circuitry.

Jan 1, 1988

By P. M. Tait, N. A. Warner

"The refining of thermally produced zinc is currently performed using a highly energy intensive and arguably outmoded process, distilling zinc under atmospheric pressure at temperatures of 900°C to 1000°C. In order to produce refined zinc more profitably and with less environmental impact a new process has been proposed. This process is generic to all thermal zinc smelters and aims to match the purity of zinc from current technology (99.99%-99.999%). The objective is to perform all separation processes using only the heat contained in the smelter product stream. Vacuum conditions are extensively used lowering operating temperatures and enhancing separations. As part of this process zinc is distilled under vacuum from high concentration obtaining a refined product with respect to lead content. A mathematical model has been developed from fundamental principles to describe this stage, the distillation of zinc across a temperature difference between two circulating zinc streams in open channels under vacuum. Heat for the distillation is supplied by the liquid phase, necessitating a large recycle to maintain high operating temperatures, and keep zinc concentrations high. In determining this model and the operating conditions required, relationships between temperature, pressure and composition are considered.IntroductionCurrent technology (the Refluxer) used to refine thermally produced zinc is both expensive and highly inefficient. At present 80% of the worlds zinc is produced via the electrolytic process reflecting the poorer performance of the thermal route, the zinc-lead blast furnace and Refluxer. A new process is proposed to overcome the drawbacks of current processes. These drawbacks are both economic and environmental. The Refluxer requires 6.6 GJ of energy to refine one tonne of SHG zinc 1, other disadvantages of the Refluxer stem from high operating temperature, around 900°C, and liquid zinc being highly corrosive towards ferrous metals. The Refluxer is constructed from silicon carbide sections which join together by means of flat machined faces making liquid metal leaks inevitable. The Refluxer columns are very sensitive to thermal cycling, thermal shock and mechanical vibration, which easily lead to structural failure. To avoid thermal stress on the columns the feed rate and feed temperature need to be kept constant which means the feed baths to the columns must be able to accept solid and liquid feed as the Refluxer has to operate when the smelter is shut down."

Jan 1, 1998

By G Lane

This paper discusses the factors that contribute to the cost effective design of a concentrator. Concentrator design and layout outcomes are functions of the team (engineer’s and owner’s) participating in each particular project. The benchmark for relatively modest projects was set in the 1980s and 1990s during the ‘gold boom’ when numerous cost effective plants were designed and constructed on a lump sum basis in a very competitive market. A number of factors contribute to lack of transference of the lessons learnt in gold plant design to concentrator design including established paradigms in the design and layout of concentrators, lack of experience in cost effective design, operator’s preferences for flowsheet and layout and simple lack of appreciation of the impact of plant footprint on materials quantities and resultant capital costs.Experiences with recent copper concentrator projects (both small and large) are used as case studies.

Aug 8, 2011

By Phil Wilkinson, Sam Whitaker

"The installation of large (XL) diameter monopiles offshore typically requires expensive and specialist plant, equipment and vessels. Due to the increasing size (weight, diameter and length) of the foundations, in order to mitigate installation risk, the use of large diameter drills is becoming an increasing requirement.Project developers often try to avoid drilling operations due to the time associated with the drilling process. This restricts projects in terms of location and it also overlooks the additional engineering benefits drilling can bring to monopile foundation design e.g. reduced foundation material and better fatigue design life.By comparing the lessons learned of three milestone overwater drilling projects; methods of maximising drill speed and decreasing drilling process time is established. Particular emphasis is put on pile internal relief drilling in a process known as Drive-Drill-Drive (DDD).The milestone overwater drilling projects are North Hoyle Offshore Wind Farm (OWF) (2003); Flamanville EPR cooling water outlet shaft (2008); and Gwynt-y-Mor OWF (2014).By evaluating the effect of foundation pile diameter on drilling process times, example foundation installation costs are derived and used to show that incorporating drilling as part of a process may not be cost prohibitive to projectsContinued advances in drill technology, and drilling rates of penetration, combined with an integrated approach to foundation design, challenge whether hammering only is always the optimal solution for XL monopile installation."

Jan 1, 2017

By Richard D. Luark, Fernanda S. Madrona

Deep foundation construction in highly developed urban settings pose unique challenges of working with deep excavations in a relatively confined area while preserving the integrity of nearby structures. For construction of the below grade parking for two, 40-story residential high-rise towers, shoring walls were required to support the maximum 70-foot-deep vertical excavation in downtown Seattle, Washington. The design of the excavation support system consisted of soil nails combined with closely spaced vertical elements to allow the top row of soil nails to be installed up to 10 feet below the ground surface avoiding conflicts with adjacent underground utilities. To limit deflections next to a seven-story building, the upper rows of soil nails were post tensioned. An instrumentation program was implemented to monitor the deflections in the shoring walls by means of vertical inclinometers and optical survey monitoring. The shoring wall performance, designed to limit deflection to less than 1-inch to meet the City of Seattle requirements, was validated by the instrumentation program. This paper provides a discussion of the design and construction of the shoring walls as well as the monitoring data collected from the instrumentation system. PROJECT INFORMATION The project consisted of two residential high-rise towers, with five levels of below grade parking, located at 707 Terry Avenue Seattle, Washington. The project site covers a half city block, approximately 120 feet by 240 feet, and is bounded by Terry Avenue to the east, Columbia Street to the north, Cherry Street to the south, and an alley to the West. Prior to the development, site was occupied by surface parking. A 7-story concrete building with two stories of below grade parking, owned by the Archdiocese of Seattle is located just west of the 16-foot-wide alley. Numerous below grade utilities consisting of electrical, telecommunications, gas, water, sanitary sewer, steam, and storm drains, surround the project. The perimeter below grade parking walls extended within inches of the property boundaries. Topographic relief across the site was on the order of 8 feet. Figure 1 illustrates the site plan view.

Jan 1, 2019

By John Chadwick

In the drive for faster development advance rates, railbound systems can be advantageous, in the right application. Skanska BS, the contractor to mine-owner VSK Mining, is driving an adit access deep into the Volouské Urchy Mountain in Slovakia to access the Gemerská Poloma talc deposit, some 70 km (43 miles) west of Ko?ice, the economic and cultural center of eastern Slovakia. Skanska began driving the Stolna Elisabeth drift in February 2007, extending 25 kg/m (50 lb/yd) rails (with 750 mm or 30 in. between sleepers) into the mountain to provide the main haulage to bring talc ore and waste out of the new mine once it gets into production. It will also be used for personnel transport, and is now almost complete. As a second egress and for ventilation, a 2.5-m- (8.2-ft-) diameter shaft of about 300-m (984-ft) depth will be sunk from surface down into the orebody. Stolna Elisabeth is a 4.2-km- (2.6-mile-) long drift mined at a cross section of some 11 m2 (118 sq ft). The finished profile is 3.2-m- (10.5-ft-) wide by 3.6-m- (11.8-ft-) high. It has been developed on a very slight incline so that water runs down it to the portal. The portal is 450 m (1,476 ft) above sea level and the end of the tunnel is 475 m (1,558 ft) above sea level. Much of the rock through which it has been developed is magnesite with a compressive strength of 150 MPA. The drilling jumbo is an Atlas Copco twin-boom Boomer 282 Raildrill drilling 42- to 45-mm- (1.6- to 1.7-in.-) diameter holes to a maximum length of 2.5 m (8.2 ft), depending on ground conditions. This is a drill jumbo that Skanska has used successfully on a number of projects during its life. The condition of the rock through which this drift has been driven is extremely variable. Five classifications of A1 and A2 and B1, B2 and [ ] B3, where A1 is the most competent rock and B3 is the worst, have been established. The amount of support used, of which more later is determined by the classification of ground conditions. At the time of the author?s visit, the project it was about 40 days ahead of schedule, thanks to high advance rates. For instance 6 m/day (20 ft/d) was the plan for the advance to be achieved in B1 rock, but sometimes the crews have achieved 8 m/day (26 ft/d) ? which is the average for the whole project ? in those conditions. In A1 the best advance rate achieved is 12 m/day (39 ft/d). The loading unit is a Häggloader 9HR-B on crawler tracks with a pony track to run on the rails. This allows the unit to be towed up and down the drift along the rails. To load from the face, the Häggloader lowers the foot of its onboard conveyor system to the floor and the backhoe pushes material on to it. This digging arm-conveyor combination produces uninterrupted, optimal volume loading of the haulage unit, without spill ? a technique offering load capacities as high as 180 m3/h (6,356 cu ft/hour). Further-more, as an electrically powered machine, it offers great

Jan 1, 2009

By V. C. Rogers, K. K. Nielson

INTRODUCTION The potential health impact from uranium milling operations is mainly associated with long-term releases of radioactive contaminants from the mill tailings.(1,2) The major mechanisms for mitigating these potential releases focus on increasing the tailings containment with the addition of migration barriers such as thick earthen covers and clay liners. Some limited investigation has also focused on reducing the radionuclide source terms. This alternative approach has some desirable features, but stringent cost requirements are placed upon source removal methods in order for them to be economically favorable. A cost effectiveness evaluation is presented herein, in which costs for containment methods are used to establish maximum cost guidelines for the source removal methods. The 230U decay chain, shown in Figure 1 contains the radioisotopes of major concern, 230Th, 226Ra, 222Rn and its short-lived alpha-emitting daughters, When the uranium is removed in the milling process, these isotopes, which are normally not removed remain in the tailings solids. The general pathways to the environment from uranium mill tailings are depicted in Figure 2. The dominant pathways are radon to the atmosphere and radium and thorium to a groundwater aquifer. The magnitude of the contaminants released into these pathways can be characterized by the following parameters: 1. The 230Th and 226Ra content. 2. The emanating power for radon. The fraction of radon generated that is free to migrate. 3. The leach rates. Source reduction methods seek to reduce the magnitude of any or all of these parameters. In general the dose rates are proportional to the source terms. For radon production the source term is proportional to the product of the radium concentration, R, and the emanating power, E, so that the risk is reduced by the same factor that reduces RE. For example, if the radium content is reduced by 95 percent, but E is increased by a factor of two, then the dose rates are reduced by an order of magnitude. The source term for the groundwater pathway is not as easy to identify. The peak in the radium concentrations in the accessable water body can also

Jan 1, 1981

By A Misra

In pursuit of solving the development rate problem the coal industry has investigated various development systems. While all the systems may improve the development rate they may not necessarily he cost effective. This paper carries out an economic analysis by developing cost models for present and future development systems for retreating longwall faces with particular reference to Australian longwalls. A critical cost analysis of various development systems is presented by taking into consideration all the costs associated with development systems. The costs have been divided into equipment, face and panel services, belt and structure, wages, roof and rib support, ventilation, consumables, power and maintenance. The annualised cost of various equipment has been obtained on the basis of depreciation rates prescribed by the Australian Taxation Office.

Jan 11, 1996

By T. Hayes

Track-type tractors remain one of the most cost-efficient stripping tools available to coal miners. Whether they are "sweetening up" a dragline operation with pre-benching or serving as the primary stripping tool, dozers can play an important role in driving down mining costs. This paper reviews the application zone of track-type tractors. In addition, it provides tips on site set-up, equipping of tractors, and machine operation which the audience can take home and apply.

Jan 1, 1996

By A. J. MacKellar

The economic viability of any project is generally assessed on an evaluation of the total capital constructed costs along with operating costs, owner's costs and market conditions. This paper will be directed to the role that Cost Engineering plays in the development of the capital costs and the cost control techniques applied thereto.

Jan 1, 1975

By David C. Slack

In recent years public concern over the environment has grown enormously. A very aware populace has exerted pressure on its legislators to enact enforcement laws dealing with pollution. Among the different types of pollution, sulfur dioxide in the air is a large and readily identifiable constituent. Air pollution, according to the EPA, costs the U, S. billions annually with SO2 a major portion of this cost. Fossil fuel and oil fired power plants release better than 50% of the SO2 in the country. Thus great attention is being focused on the methods power plants are utilizing to control emissions. With a time schedule of compliance by early 1975, decisions are being made and systems chosen that will have to perform - and perform - and perform. Therefore, a large majority of the power plants are presently considering lime in the planning and operational phases of their stack gas desulfurization programs. The reasons for this are: Lime scrubbing has more development and pilot work completed than most other methods; the removal potentials are relatively high; the stoichiometric ratios are not excessive and, the process has relatively low capital and operating costs. Many pilot plant studies have been completed recently using lime and the first operational full scale lime scrubbing unit is now in the start-up phase at this time at Duquesne Light Company's Phillips Station here in Pittsburgh. Also, because of completed development work, the problem of solid waste disposal will also be tackled at this power station at the same time. There are various methods for the use of lime in scrubbing for S02, but for our purposes I would like to describe the cyclic wet lime process. This method called "Scrubber Addition of Lime" involves contacting the flue gas with a slurry of milk of lime.

Jan 1, 1973

By A. J. Richardson

Before the development of the underground stoping and mining costs can be considered, certain facts about the ore body, the proposed mine, markets, etc., must be known or determined. In the case to be studied, the zinc-lead mineralization occurred with a narrow vertically dipping structure of undetermined length and vertical extent. Exploration completed to date has revealed 6.5 mil¬lion st t of proven reserves. A further 820,000 st of in¬dicated reserves has been outlined and this tonnage is considered capable of being expanded by a factor of approximately four after more detailed drilling. After studying the market conditions and completing a very preliminary feasibility study, it was decided that production would be 730,000 stpy (or 2000 stpd) of ore. First year production would be at the rate of 1500 stpd. The main design criteria for the selection of the min¬ing methods are minimizing surface subsidence, maxi¬mum recovery of the ore body, maximum degree of grade control, maximum productivity, and safe working conditions. Two basic extraction systems are considered capable of meeting these requirements: mechanized cut¬and-fill stoping and sublevel long-hole stoping with filling. The primary development system of the mine has been designed to give maximum flexibility in stoping systems and layout and to permit changes if considered necessary as a consequence of actual production ex¬perience. At the present time, access to the mine is by a circu¬lar concrete lined vertical shaft, 16 ft diam, sunk to a depth of 1380 ft. Two exploration levels have been driven within the ore zone at depths of 165 and 1246 ft below the surface outcrop. The development to date had the objective of sampling the mineralization and produc¬ing detailed information on the outline of the ore body and the distribution and controls of zinc and lead values. In an attempt to satisfy the basic design criteria for the mine, it was decided that production would be best achieved by a combination of 40% sublevel long-hole stoping and 60% cut-and-fill mining. Costs of exploration and capital development of per¬manent underground facilities are normally written off over the life of a mine. Production expenditures, on the other hand, are of a temporary nature and are normally charged as and when incurred as an operating expense. Reasonably accurate predictions of mine production costs can be built up from engineering design and estimates of individual mine activities for ultimate inclusion in the comprehen¬sive data required for financial decision making. The simulated operations can be costed on a detailed basis in the form of a monthly operating budget. The budget format can be generalized or detailed, depending upon the scope of the project. However, ex¬perience suggests that a fairly detailed format has the advantage of assuring that all significant cost items are included. For underground costing it is suggested that the budget structure include five major cost centers (i.e. development, diamond drilling, ore extraction, hoisting/ transportation, and general mine expense). These in turn are detailed under numerous subheadings. The mechanism for compiling an operating budget will be illustrated. Because of its relative simplicity, ore extraction under sublevel long-hole stoping has been chosen for illustration. All other activities, simple or complex, can be estimated in similar fashion. BLOCK AND STOPE DEVELOPMENT Long-hole blocks, used where advantageous, will be up to 250 ft in height, depending upon the vertical con¬tinuity of the mineralization, and approximately 300 ft long. Drawpoints will be at 36-ft intervals and serviced by loading crosscuts driven from a footwall drift parallel to and close to the ore zone. Pillars between the stopes will be 50 ft wide. Stopes will be drilled off with vertical rings of blastholes drilled from sublevels approximately 60 ft apart vertically. This drilling will be done by percussion drilling machines (31/2 in.) mounted on a trackless drilling rig. Load¬haul-dump (LHD) equipment will be used to move broken ore from the drawpoints to the orepass connecting to rail haulage systems. On completion, long-hole stopes will be backfilled to prevent caving and to facili¬tate later pillar removal. From a planned stope layout, a forecast of produc¬tion and development is made in Table 1. Table 1. Block Tonnage and Stope Development Quantity Ore Waste Total ore block 375,000 st 2 stopes 310,000 st 1 pillar 65,000 st Access crosscuts, 4 at 100 ft 400 ft Drill sublevel drifts, 6 at 300 ft 1800 ft Stope raises, 3 at 250 ft 750 ft Undercut sublevel drifts, 2 at 300 ft 600 ft Loadout crosscuts at 35-ft intervals 550 ft 100 ft 3300 ft 500 ft Total development footage 3800 ft Tons per ft of development 987 st

Jan 1, 1982

By James E. Breland

Society's growing concern over the number of deaths and serious injuries occurring in coal mines, coupled with a growing sentiment to develop and promulgate improved mandatory health or safety standards, was no doubt the basis for passage by Congress of the Federal Coal Mine Health and Safety Act of 1969, being Public Law 91-173 approved December 30, 1969. Even prior to passage of the Act, strong reaction was quick to arise concerning the effect of such legislation upon the coal mining industry. Initial response varied from mild forms of criticism to' the expressed fear of complete suspension of operations and mine closings. Primarily responsible for these early forms of skepticism were speculative or fanciful cost estimates which led operators to contemplate a less than prosperous existence. However, in spite of this unfavorable outlook, the majority of the operators have faced up to the situation and have been striving to comply with the Act and the regulations issued thereunder. While encouraging strides have been made toward implementation of the Act, it must be recognized that noticeable results have been attained only through many months of determined effort and, as anticipated, at quite heavy expense to the operator.

Jan 1, 1973

By B King

Processing expansion options and mining fleet purchases are major cost items that receive due focus in most cost management analyses. If these items are not needed, they can significantly reduce a project’s short-term capital spend and improve profitability and value to shareholders. When conditions change prior to purchasing equipment, the purchase may no longer represent the best decision for the project.This paper outlines a structured approach for making major cost decisions to improve the value of a project, determining which items should be kept and which should be eliminated.CITATION:King, B, 2014. Cost management optimisation, in Proceedings Orebody Modelling and Strategic Mine Planning Symposium 2014 , pp 7–12 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 24, 2014

By V. O. Tenorio, A. M. Sanzana

"INTRODUCTION This report summarizes a research about mines which are transitioning from surface to underground mines. It is a guide to make strategic decision in these increasing amount of mines. Capital and operating costs are linked with major strategic decisions in the exploitation of an ore body. This is the reason to analyze costs related to this transitioning world trend. There are multiple reasons for transitioning from Open-Pit (OP) to Underground (UG) mining. The primary economic incentive for going UG is that the cost per unit of ore from OP mining eventually exceeds that from UG extraction. This is due to the excessive amount of waste that must be extracted when the depth of the pit increases. In addition, the reduced footprint of UG operations contributes to diminish the environmental impact on the surface. Furthermore, it becomes more difficult to find new, easily accessible ore deposits. Then, a cost model helps to guide strategic decisions in the transition from OP-to-UG mines. Objectives • Propose a methodology to guide strategic decisions in the transition from OP to UG mining. • Create an integrated cost model based on existing OP and UG individual cases. • Identify and characterize a group of mines transitioning from OP to UG. • Analyze operating and capital costs of a particular transitioning from OP to UG mine. MAIN CONCEPTS Open Pit Mining (OP) The main decisions in OP mines are: Cut-off grade, Stripping Ratio (SR), equipment size, stockpiling, pushback sequencing and processing capacity. The stripping ratio is measured as the tonnage of waste required to be removed to reach and extract one tonne of ore. The material extracted from an OP exploitation can be classified as waste and ore. This classification depends on the cut-off grade. The cut-off grade indicates the minimum amount of material with economic value in a block. The waste from the mine is going to the waste dump and the ore is going to stockpiles where will be store and then going to the processing plant (leaching or concentrator). The waste of this process is going to the tailing pond big pools. And the product is sell in the market. Another important issue is the mining capacity which is going to define the investment in mining equipment. Underground Mining (UG) The first decision to be made in order to open an underground mine is selecting the mining method. Caving methods are the more productive and have lower operating costs. There are three caving methods: Block Caving (BC), Panel Caving (BC) and Sublevel Caving (BC). The access development could be an horizontal adit or a straight or downspiral decline. A shaft could be used for inclined or vertical access. Other key issues are: equipment capacity, cut-off grade and production sequencing."

Jan 1, 2016

By S. A., Taylor

IT is very difficult to arrive at exact figures for the cost of maintaining excess capacity of coal mines, but we can approximate the various items. To do this, I will take the Pittsburgh district of Pennsylvania and the other districts can be compared to this one. I believe we can take the depreciation item of plant ac- count as a whole and discount it over a period of 20 years; while this is high for some items it is low for others and I believe it is not far from a safe basis. This equals 5 per cent per year, and this depreciation will take place about as rapidly when not in use as when in operation. A fair value in the Pittsburgh district to- day for plant and houses is close to $4 per ton per annum. At 5 per cent this will equal 20 cents per ton per year. The item of keeping the mine in condition, free from water, etc., will be close to an average of 2y2 cents per ton per annum.

Jan 1, 1928

By L. B. Rothfeld

This paper presents preliminary results of a Bureau of Mines study of potential costs of compliance with the federal Resource Conservation and Recovery Act for copper heap and dump leaching. Specifically, estimates are given for the cost of compliance with possible technology standards that would require the use of liners and closure caps regardless of terrain and underlying geology and hydrology. This cost is compared to the co lit of compliance with present state requirements for terrain a typical of those found at existing copper leach sites in the Southwestern U.S.A. Incremental compliance costs are found to widely exceed screening criteria used by the Environmental Protection Agency to denote significant economic impact.

Jan 1, 1993

By L. B. Rothfeld

In anticipation of RCRA reauthorization and subsequent rulemakings, the Bureau of Mines is studying potential economic impacts of RCRA on mining. For U.S. copper heap and dump leaching operations in a variety of terrains, the cost of liners and closure caps to comply with RCRA Subtitle D ranges from $55 million to $158 million per site, or about $500,000 per hectare ($200,000 per acre). If RCRA technology requirements and financial assurance had been required on current operations, 27 to 100% of production from heap and dump leaching would have been non-competitive at market prices between $2,310 and $1,650 per metric ton ($1.047 and $0.75 per pound) copper. This result is based on an analysis that considers return on capital, corporate overhead, marketing, and freight costs as well as operating and maintenance costs at the mine. Over the same price range, economic analysis using the IMPLAN input-output model on one-or two-county regions indicates potential net job losses of about 570 to 2,700 jobs. Total economic output of those counties could be reduced by about $180 million to $360 million.

Jan 1, 1994