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By C. E. Sims

THE following presentation is more a summation of observations made and conclusions drawn over a period of three to four years than a report of experiments made with a definite objective. Tests were made, of course, to verify hypotheses, but the large mass of data was collected incidental to other work. Lack of space forbids the presentation of more than a few of those data. On the other hand, it is unfortunate that some of the most illustrative examples are not available. The work is admit¬tedly incomplete but is presented in the hope that it will stimulate further work along a similar line. The answers to many of the questions left unanswered by this paper will be found only after long and painstaking study, such as is scarcely possible in a commercial laboratory where problems of immediate importance consume most of the time. It is also hoped that this study will in some measure throw light on some of the dark spots of steelmaking, and help to clarify some of the annoyingly vague references to "good practice." Only cast steel was studied and the conclusions apply to cast steel, but it is the authors' opinion that the quality of rolled or forged steel is largely dependent on the quality of the ingot, which is a casting. Although this discussion deals almost entirely with inclusions, it is important that the reader understand that only those natural inclusions which owe their origin to the solubility of certain materials in steel and to the chemical reactions of steelmaking are considered. Accidental inclusions, such as entrapped particles of slag, nozzle, runner, or mold are outside its province and are not considered. Such accidental inclusions are always dangerous and should be carefully avoided.

Jan 1, 1932

By A. J. Van Der Merwe, N. A. Barcza, M. Boyes, N. R. Clarke

Oriel Resources has with the support of SRK (Cardiff) and DRA and Mintek (South Africa) carried our a Definitive Feasibility Study (DFS) to recover high grade chrome ore from its Voskhod deposit very close to Chromtau that is near to Aktobe in Kazakhstan. Voskhod is a large podiform ore body that contains nearly 20million tons of chromite grading about 48% Cr2O3 and with a Cr/Fe ratio of over 3/1. Selected bore hole core samples were obtained from the confirmation drilling programme during the DFS phase of the project in 2005 for pilot plant test work to design the mineral beneficiation plant. The chrome ore is characterised by four main types, namely high grade lumpy and chip sized ore, high-grade fine ore, lower grade (subordinate) lumpy ore and lower grade fine ore. Gravity separation tests carried out by Mintek showed that the large density difference between the high grade (at over 3.7t/m3) and lower grade ore (at less than 2.97t/m3 ) should ensure very good separation in the DMS drum and DMS cyclone stages resulting in high recoveries of Cr2O3 - rich chromite ore products. Very high chrome ore grades of up to 64% Cr2O3 and Cr/Fe ratios of 4 to 1 were obtained at an 80% mass yield from selected samples of lumpy and chip ore size fractions. The overall predicted recovery from the design work on the mine and plant is expected to be closer to about 70% of the ROM (Run of Mine) feed. DRA has completed the design of the mineral processing plant to recover about 900 kt/a of chrome ore from the 1.3 Mt/a of ROM feed. The plant has been sized to handle up to 1.5 Mt/a and will also be able to address the variability in ore feed as different regions in the deposit are mined. The lumpy high-grade product and the chips size fraction is expected to grade around 48% Cr2O3 and the fine concentrate between 56 and 58%Cr2O3 The ratio of products is expected to be about 70, 15 and 15 percent respectively but will vary to some extent during the expected 20 year life of the mining operation based on including the associated adjacent Karaagash deposit together with the Voskhod deposit. The prediction of the product split from the processing plant depends on many factors such as the geology, mineralogy, mining, handling and metallurgical processing. An overall process flow model has been developed that covers the ROM ore, the various types of chrome ore feed to the mineral beneficiation plant, the projected products and the subsequent smelting of the chromite ores to produce High Carbon Ferrochromium. The model allows the optimisation of the quality of the chrome ore in terms of Cr2O3 content, Cr/Fe ratio and recovery to be studied. This information can also be evaluated in terms of the economic requirements for smelting High Carbon Ferrochromium (HCFeCr) at a given location. Oriel is planning to supply chrome ore to Ferrochrome plants mostly in China and Russia. The use of the model will assist to maximise the value of the Voskhod project to its owners and also assist the users of the chrome ore in selecting the appropriate supply quality to meet their needs for a given smelter operation. The smelting of high grade Voskhod ore requires a limited addition of about 5 to 7% quartz to flux the MgO and Al2O3 and about 25% coke as a reducing agent to produce an alloy containing from about 65 to 74%chromium depending on the Cr/Fe ratio.

Jan 1, 2007

By R. Ward Stahl

Fire is still a scourge of the mining industry, although many precautionary measures have been and are being taken to reduce the incidence of fires and to combat them. Generally, the success of fire extinguishment is directly proportional to the speed of attack, and the speed of attack in turn is dependent upon the availability of materials and trained personnel. Preparations for firefighting are frequently neglected because they are considered emergency needs and do not contribute directly to coal production. Fire incites panic--a fact often overlooked--but such chaos can be minimized by training people in the use of equipment and procedures. Heat, smoke, and fumes hamper the approach to a fire; however, a relatively new technique of attacking fires with high-expansion foam may well contribute greatly to firefighting. The foam is transported to a fire and used at a considerable distance, thus permitting the firefighters to work more safely.

Jan 1, 1962

By Craig Wilbour, John Stimberis, Rob Gibson, Lee Redden

An avalanche is a snow slide. A simple explanation is that the snow on a slope will slide (avalanche) when the snow strength can no longer support its own weight. Snow avalanches happen when the load from snowfall or ram on a snow slope increases faster than the stre n g t h of the bonds between snow crystals. Avalanches can also happen when the strength of a buried layer decre a s e s . Active avalanche control is the intentional triggering of slides. Washington State Department of Tr a n s p o r t a t i o n (WSDOT), does avalanche control to minimize the chance of natural avalanches into traffic. Avalanche control work is done all over the world by: highway departments, ski resorts, railroads, mining operations, utility companies, and other activities threatened by avalanches. The majority of active avalanche control is done with skis, by explosives, or exploding projectiles. Blasting techniques and practices have evolved to fit the specific needs of avalanche contro l . Passive avalanche control is the use of engineered s t r u c t u res such as snow sheds, bridges, or snow supporting devices, to permanently protect a specific location from avalanches. The problem is that these struct u res are extremely expensive. There are a variety of passive avalanche control measures as well as the active m e a s u res used to protect the highway over Snoqualmie Pass. As traffic volume and use expectations incre a s e , m o re extensive passive measures will be needed to keep avalanches from affecting the highway.

Jan 1, 2004

Mine ore, reduced to minus 71/2 in. by primary crushers, is stockpiled. Size reduction of the reclaimed material by secondary and tertiary crushers is achieved in a closed circuit crushing and screening plant. Screen products of minus 30 mm plus 6 mm lump, and minus 6 mm fines are stockpiled by conveyor and stacker systems prior to rail despatch 10 port facilities at Dampier. Crusher settings, determined from experience, are closely maintained to obtain the desired lump/fines ratio. Chemical and sizing analyses from automatically taken samples are used to maintain product specifications. The beneficiation plant upgrades the fines product when this is necessary. The recorded wear life of individual plant components is the basis of a planned maintenance programme.

Jan 1, 1970

By Terry McMahon

This Bureau of Mines report describes the rock burst problem and measures being taken for its alleviation in the deep underground mines of the Coeur d?Alene Mining District in northern Idaho. The geologic features that contribute to rock bursting in that area are discussed and briefly compared with similar features in other burst-prone mining districts throughout the world. Early investigations to understand and control rock bursting are reviewed, and the mechanics of bursting are discussed in terms of the seismic events that produce them. Sources of strain energy, rock failure mechanisms, and the conditions necessary for the release of seismic energy are covered in detail. Rock burst research in the Coeur d'Alene District is described. Included are discussions of the microseismic method as both a research tool and as a rock burst monitoring system, rock destress blasting and preconditioning methods for rock burst control, and stoping methods currently being developed to reduce rock burst hazards.

Jan 1, 1988

By S. J. Parry, D. M. Evans, D. L. Buchanan

The reef is a PGE-enriched sulphidic zone, stratigraphically associated with and overlapping the C1d chromitite layer, that was formed by magmatic processes at the 1e-1d subunit boundary. Two distinct patterns of PGE enrichment are related to the presence or absence of Cr2O3 oversaturation during sulphide liquation that occurred during a complex sequence of changes at this boundary

Apr 1, 1996

By E. G. Davis

As a result of interest by the U.S. Department of Energy, the Bureau of Mines conducted studies on comminution of bituminous coal to the 2-um particle size range using the Bureau-developed turbomill. Both "plant grind" size (80 pct minus 75 um) and microsize (minus 10 um) coal can be produced in a single step with the Bureau's turbomill. In 15 min, minus 2.4-mm coal milled in water with steel shot was reduced tc 65 pct minus 75 m and 26 pct minus 2 m, with an energy requirement of 139 kW^h/mt coal. Plant-grind size coal milled in water, with Ottawa sand as the milling medium, was reduced to more than 45 pct minus 2 m in 15 min, at 175 kW?h/mt. Plant-grind coal milled with steel shot was reduced to 57 pct minus 2 um in 15 min, at 138 kW?h/mt. When diesel fuel was substituted for water, the milling of the coal was less effective and less energy efficient. Keeping the pulp dispersed was critical. Without an effective dispersant, the slurry became very viscous with increased milling time, resulting in higher energy requirements, low heat dissipation, and poor grinding efficiency.

Jan 1, 1987

By W. Tom Witherspoon

Many Engineers are skeptical and cautious about skin friction factors for the portion of drilled piers where temporary casing is used. This has led Texas engineers in many cases to eliminate skin friction values for the cased section, which necessitates deepening the drilled shaft to compensate for the loss in the cased section. This paper has provided field test data that supports inclusion of skin friction in the temporarily cased hole section as a contributing factor in total axial capacity as comparable to open hole shafts of the same diameter. Reasons for the resultant friction values are provided including; remolding of the cased clay section and the increased pier circumference necessitated by the oversized casing required for adherence to the engineers shaft diameter specification.

Jan 1, 2009

By Robert Thomas

DURING the past 20 years the advantages of reinforced concrete as a substitute for timbering in so-called permanent mine openings have been fully recognized, and its use has become almost general practice in the larger mines. The term "permanent mine openings" is used advisedly and is assumed to embrace such openings as ventilation shafts, material shafts and main hoisting shafts, with their adjacent stations, ore pockets, etc. In the concreting of such openings, the actual cost as compared with that of timbering has been of secondary consideration, the support of large openings, elimination of fire hazards, and prevention of inter-ruptions to production being the primary objectives. The ultimate economy, however, of lining such openings with concrete when they are to remain in use during the entire life of the mine, and especially when they are in bad ground, will seldom be questioned. During this 20-year period there have appeared in engineering publi-cations a number of interesting articles on the use of concrete in mines, nearly always describing some specific project of major proportions, such as the concreting of ore pockets, large stations or main hoisting shafts. Such articles are unquestionably of interest, since the projects require considerable engineering thought, not only in the method of excavating large openings, but in their engineering design as well. Concreting of main hoisting shafts, particularly when this is done with the least possible interruption to production, provides some interesting problems. The concreting of haulage drifts of the dimensions commonly found in mines, on the other hand, soon develops into a routine operation, there-fore it is with some hesitancy that the author attempts to prepare an article on this subject. The writer can recall only one article on the concreting of mine haulageways, although several articles have been written concerning the concreting of tunnels considerably larger in dimensions than those required in general mine practice.

Jan 1, 1936

By L. Douglas Smoot

This research program was a theoretical and experimental study to improve understanding of the mechanisms of flame extinguishment in methane-coal dust-air-suppressant mixtures. This final report is a summary of work conducted during this grant, sponsored by the Bureau of Mines. Work under this new grant closely followed a recently completed contract for the Bureau of Mines that was conducted at this laboratory, and results of the earlier contract are also summarized in this report, as required in the grant work statement. The experimental effort was concentrated on supplementing the limited data available to describe laminar flames in methane-coal-air-suppressant mixtures. The theoretical effort was devoted to developing a comprehensive model that could be used to explain and extrapolate the experimental data. Selected for study were the following flames: Methane-air; methane-air plus KHCO , KNO3, K C204, KBr, Al203 or Monnex; methane-coal dust-air; coal dust-air with size, volatiles content, and concentration as variables, and coal dust-air plus KHCO3, KN03, KBr, NaCl or CaC03.

Jan 1, 1978

By Thomas A. Phillips

The Bureau of Mines presents an economic evaluation of its leach- electrolysis process for recovering lead from galena concentrate. In this process, galena concentrate is leached at 200° F with a ferric chloride-brine solution to extract lead while leaving sulfur in the residue, which is discarded. The solution after leaching is concentrated and cooled to 100° F by vacuum evaporation to crystallize lead chloride from which metallic lead and chlorine are produced by fused-salt electrolysis. Chlorine is used to regenerate the spent leach solution before it is recycled. A cost estimate was prepared for a plant capable of producing 300 tons per day of lead, 330 days per year. The fixed capital cost on a third-quarter 1977 basis Fs $27,500,000, and the estimated operating cost without a charge for the concentrate is 4.9 cents per pound of lead product. Assuming a selling price for corroding-grade lead of 31 cents per pound and a charge for the concentrate of $332 per ton, the interest rate of return on investment after taxes is 16 percent. The leach-electrolysis process, therefore, appears to be economically attractive under the assumed market conditions.

Jan 1, 1978

By Eric Konigsmann, Doug Spence, Scott Jones

"The Williams Gold Mine is located in the Hemlo camp of Northwestern Ontario and is one of Canada’s largest gold producers with annual gold production of 400,000 ounces. The operation is owned and operated as a 50:50 joint venture between Teck Cominco Limited and Barrick Gold Corporation.In July 2002, the decision was made to introduce paste backfill produced from mill tailings as the primary backfilling method in the mine and to decommission the existing cemented rock fill system. This required a fast track design to be completed for both a surface paste backfill plant and an underground paste distribution system followed by an intensive construction phase through the fall and winter of 2002/2003. The paste system was commissioned on schedule in late April 2003 and was in full production within 4 weeks of the initial start up. Overall, the project duration was only 42 weeks from initial project release to commissioning and it was completed $2 million under the budget of $14 million.This paper presents an overview of the lab testwork and process flowsheet design, process control methodology and describes the construction and commissioning phases of the project. Preliminary paste plant performance and operating cost data is also presented.IntroductionThe Williams Gold Mine is located in Northwestern Ontario adjacent to the Trans Canada Highway in the Hemlo Gold Camp as shown in figure 1. The mine currently produces approximately 400,000 ounces of gold annually and is owned and operated as a 50:50 joint venture between Teck Cominco and Barrick.The deposit was discovered as part of the Hemlo gold discovery in 1981 which led to the development of three mining operations within a 2 kilometre radius. The neighboring operations are the Golden Giant Mine, currently owned and operated by Newmont Gold, and the David Bell Mine, currently owned and operated by the same joint venture which is running Williams. David Bell ore is blended and milled in the Williams concentrator."

Jan 1, 2004

By R. R. Faddick

Properties of coal-water slurries were studied in two separate phases: a rheological phase and a pipeline phase. The purpose of the study was to investigate the rheology of fine coal slurries to determine the effect of fine particles in a coal-water mixture on the economic and operational aspects of a slurry pipeline. The slurries consisted of 6x10, 100x200, and 200x0 mesh coal in water. A fourth slurry consisted of a mixture of the 6x10 and 200x0 mesh coal. Brookfield and Stormer rotational viscometers were used to obtain or simulate rheograms for a range of fine coal concentrations, temperatures, and particle sizes. All the slurries were pumped through a 1-inch diameter acrylic pipe. The fine coal slurries were found to encompass four distinct rheological models. The narrow size distribution of the coal slurries was apparently responsible for the high pipewall friction losses, more so than the particle size. A scale-up method applicable to Bingham plastics was found satisfactory for predicting pipeline headlosses of fine coal slurries from rheology data.

Jan 1, 1974

By Eric R. Bauer

The U.S. Bureau of Mines, in cooperation with a southwestern Pennsylvania coal mining company, BethEnergy Inc., Mine No. 60, Eighty-Four, PA, recently assessed the feasibility of using pre driven recovery rooms (open entry recovery) when longwall mining to increase the productivity of the coal extraction process. Two recovery entries, one 200 ft long and another 600 ft long, were supplementally supported with bolts, channels, and wire mesh, and either fly ash-cement piers or fiber cribs, then mined into by the longwall face. The recovery entries showed little deterioration as the supports provided adequate resistance to the front abutment loading. Compared with conventional recovery methods, the fullface recovery room allowed for panel extraction to be completed nearly 17 days faster and to reduce face equipment move time by 1-1/2 days. The potential productivity increase accompanying recovery rooms is greater than the cost of supporting a fullface recovery room. The fullface recovery room provided over 18 additional production days and the opportunity to mine nearly 68,400 st of coal per year per panel mined.

Jan 1, 1989

By T. M. Tharp

Attempts to analyze the stability of slopes, foundations and underground openings in discontinuously jointed rock have generally assumed full joint continuity or ignored the role of stress concentration at the tips of discontinuous joints. The most common approach in rock slopes is to consider a failure plane consisting of coplanar, but discontinuous joints separated by intact rock. The intact rock is assumed to fail in shear as a MohrCoulomb material and to contribute a strength component proportional to the fraction of the failure plane that is intact rock (Jennings, 1970). An alternative approach by Stimpson (1978) applies Lajtai's (1969) theory for shear strength of discontinuously jointed rock masses to slopes. Lajtai allows tensile failure in the intact rock of a composite surface. This theory yields lower shear strength than that of Jennings (1970) except in the presence of high normal compression, however it also fails to account for the stress concentration at the tip of a discontinuous joint. For a given slope geometry, stresses increase with slope height, but in both of these models rock mass strength is independent of size. A fracture mechanics analysis shows that particularly for small slopes, rock mass strength decreases with size.A continuous stepped failure surface produced by propagation of preexisting joints may develop at lower shear stress than pre¬dicted by the Jennings (1970) or Lajtai (1969) models. This fracture mechanics model applied to slopes in competent rock is more realistic in its assumed failure mechanism, and for large slopes more conservative than the Jennings (1970) or Lajtai models (1969).

Jan 1, 1984

By S. Yalcin, S. Chang, M. Oliazadeh

"Present day ferrochrome demand is on the rise, as more steel products require this material as an alloying element. It is projected that the stainless steel demands will grow by about 6% per year to 2020; as such, the supply and demand for ferrochrome appears promising. Chromite is the only economical source of chromium, which is commonly upgraded by gravity techniques and used to produce ferrochrome. The gravity separation methods are not able to treat particles finer than 0.1 mm efficiently. Therefore, in processing chromite ore, a large quantity of fine tailings; some 25% of the mineral values have been reported as lost in slimes during processing. The chromite content of these tailings is typically 10-20% Cr2O3 and accumulates into a large stockpile/pond. A chromite concentrator project in Canada, considers a cut-off grade of 20% Cr2O3. This indicates further processing might be required to recover chromite from low-grade stockpile when the high-grade deposits are depleted. Low-grade ores normally indicates disseminated chromite ores with low liberation and consequently fine comminution. This paper reviews the techniques for treating chromite from fine material and provides some examples of successful operations in fine chromite recovery. Beneficiation techniques for fine particles include gravity separation, magnetic separation, and flotation. Several case studies are presented here that used one of the aforementioned methods or a combination of them.INTRODUCTIONChromium is one of the essential element and important material for production of ferrochrome alloys and special steel. Chromium occurs as chromium spinel, a complex mineral with chemical composition of (Fe, Mg) O (Cr, Al, Fe)2 . Chromite ore is exclusively found in ultramafic igneous rocks. Major associated gangue minerals are talc, quartz, serpentine and iron oxides.Chromite is a heavy mineral, with a specific gravity in the range of 4-5, relative to associated gangue minerals in a chromite ore, which makes gravity concentration a suitable process. This is generally true for coarse particles, ranging from 1 mm to 100 mm. Typically, the ore is prepared by crushing (and grinding if necessary) to a suitable size and the material is sized for the appropriate beneficiation techniques. It is common to use heavy medium separation (e.g. dense media drums, dense media cyclones, or Dyna Whirlpool) to treat coarse size and intermediate size particles (1 – 100 mm). Intermediate and fine particles (20 – 0.1 mm) could be treated by jigs. Spirals, shaking table, sluices, and Reichert cones can upgrade fine particles. Figure 1 shows a typical chromite beneficiation process flowsheet. As can be seen, it consists of three major sections: feed preparation (crushing, screening), de-sliming, and concentration. Beneficiation recoveries range from 80 to 85%, although the ratio of upgrading is only 2 to 3."

Jan 1, 2013

By Shail J. Butani

In 1986 the Bureau of Mines conducted a probability sample survey, Mining Industry Population Survey, to measure such employee characteristics as occupation; principal equipment operated; work location at the mine; present job, present company, and total mining experience; job-related training during the last 2 years; age; sex; race; and education. The population estimates are necessary to properly analyze the Mine Safety and Health Administration (MSHA) injury (includes illness and fatality data) statistics; that is, to compare and contrast injury rates for various subpopulations in order to identify those groups that are exhibiting higher than average injury rates. This report uses the survey's results to characterize the U.S. nonmetallic mining workforce from March through September 1986. Similar reports have been prepared for the metallic, stone, and sand and gravel mining industries, as well as for the entire metal and nonmetal mining (includes metallic, stone, sand and gravel, and nonmetallic industries) sector and the cool mining sector.

Jan 1, 1988

By Shahriar Vandani, Andy Herlache, Andrew Bro, Wei-Yu Chen

"The Presidio Parkway Project will reconstruct approximately 1.6 miles (2.6 km) of highway, including seven new bridges, three new cut-and-cover tunnels, and a new undercrossing connecting the Golden Gate Bridge to the City of San Francisco, California. The project site is approximately 5.6 miles (9 km) from the San Andreas Fault, capable of producing magnitude 8 earthquakes. Seismic hazards include strong ground shaking, soil liquefaction, ground settlement, and lateral spreading. The Main Post Tunnels (MPT) are two parallel cut-and-cover concrete tunnels approximately 1100 feet (335 m) long. The MPT site underlain by loose sandy soils susceptible to liquefaction and lateral spreading. Excessive racking and permanent displacement due to lateral spreading are identified as the main seismic vulnerabilities. Analyses showed that without ground improvement, tunnel racking of over 5 inches (13 cm) and permanent lateral tunnel displacement of up to 15 feet (4.6 m) could be anticipated. Cement Deep Soil Mixing (CDSM) is proposed to mitigate slope displacements and increase lateral stiffness of the foundation materials to reduce tunnel racking. A series of 2-D fully dynamic time domain analyses were performed to evaluate the performance of CDSM, confirm that performance criteria are met, and optimize the CDSM geometry and replacement ratio. The use of CDSM reduced the permanent tunnel displacement from over 15 feet (4.6 m) to less than 1 inch (2.5 cm). Tunnel racking was also reduced significantly. The adopted analytical approach presented herein can be considered for future tunnel design projects in seismically active areas.Introduction The Main Post Tunnels (MPT) are an integral component of the Presidio Parkway project which will replace the existing facility with a new multi-lane roadway between the Golden Gate Bridge and the City of San Francisco. The MPT are two parallel cut-and-cover concrete tunnels approximately 1100 feet (335 m) in length. As shown on Figure 1, the project site is located in a seismically active area, about 5.6 miles (9 km) east of the San Andreas Fault and 12.5 miles (20 km) west of the Hayward Fault. The San Andreas Fault can generate magnitude 8 earthquakes while the Hayward Fault is capable of producing magnitude 7.3 earthquakes. The design Peak Bedrock Acceleration (PBA) is 0.46g for the upper level design earthquake. The project alignment is shown on Figure 2."

Jan 1, 2014

By W. R. Reed, G. J. Joy

The personal dust monitor (PDM) is a new coal mine respirable dust sampling instrument that has the ability to provide accurate end-of-shift and real-time respirable dust exposure data. A hypothetical cost comparison analysis of the PDM with the current MSHA-required coal mine dust personal sampling unit (CMDPSU) was conducted. Some simplifying assumptions were made to create hypothetical cases for analysis. This analysis was limited to respirable dust sampling of one designated operator (DO) and one designated area (DA) for one mechanized mining unit (MMU) which operated one shift per day. Only sampling intervals of the minimum sampling as required by federal regulations and continuous sampling were analyzed. This cost analysis from the company perspective is only one part of an overall cost/benefit analysis that should be conducted at a later date. In this analysis, company-associated costs were examined for both sampling systems. Results show that the PDM, while it has high initial costs of approximately $12,600, may be a cost-effective sampling system for measuring coal mine respirable dust. The estimated present worth cost of conducting the minimum required dust sampling over a five-year life ranges from $5,850 to $19,000 for the CMDPSU, with the higher present worth cost including costs of potential violations incurred. For the PDM, the comparable present worth cost is $14,900 to $19,300, with the lower present worth cost representing a lower capital cost of $8,150 per unit if more than 850 units are produced per year, while the higher present worth cost reflects the result with the higher $12,600 capital cost.