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By E. D. Attanasi, J. H. DeYoung

Although fraught with problems of completeness and comparability, statistical measures of mineral exploration in the United States point downward for 1985 and 1986. Exploration expenditures in the US domestic and Canadian companies have dropped steadily from a 1981 peak of about $400 million to about $190 million in 1985, according to a recent survey of exploration statistics by the Society of Economic Geologists (SEG). These totals have been adjusted to account for differences in the estimated coverage of exploration activity by each year's SEG survey. This survey is now supported by the US Bureau of Mines and the US Geological Survey (USGS). The drop in current dollar amounts does not take into account the effects of inflation, which would make the decline even more substantial. Metals Economics Group of Boulder, CO reported in its Mine Development Bimonthly that announced exploration budgets for 1986 indicated an even larger drop for the past year. Worldwide expenditures for companies covered by the report dropped from about $800 million to $750 million from 1984 to 1985 and plummeted to about $600 million in 1986. Exploration activity has been declining since about 1981. This has been due in large part to low prices for base and ferrous metals. Exploration budgets of mining subsidiaries of oil companies have plunged as parent companies try to conserve cash in the face of the sharpest decline in oil prices in nearly 40 years. Mining subsidiaries of petroleum firms that were either liquidated or divested recently include Anaconda (Arco), Getty Mining (Texaco), and Cyprus Minerals (Amoco). Exploration expenditures in Canada increased in 1984 in regions where gold is a traditional target. But there was little change in 1985, based on preliminary statistics from a federal government report (Energy, Mines and Resources Mineral Bulletin MR 211, 1986). Claim staking in Canada dropped 23% from 1983 to 1984 and declined slightly in 1985. Diamond drilling increased in 1984 to the record levels of 1980 and 1981. The ratio of exploration to development expenditures in Canada, however, fell from a 1981 peak of 0.9 to about 0.5 in 1984. The report also indicated that, since 1981, there has been a massive write-off of reserves with high production costs. Reserve levels of all major metals were down since 1981. Precious metals exploration accounts for an increasing share of total mineral exploration. For example, the SEG data show that 90% of 1985 US exploration expenses by domestic mining companies were devoted to base and precious metals, compared with only 51% in 1980. The restructuring of the mining industry has been accompanied by a decrease in response to the last two SEG annual surveys, but the trend towards precious metals was already established in earlier years. Trade journal reports suggest that the reduction of long-term interest rates has encouraged the mining industry to make cost-reducing capital investments in existing mines to maintain competitiveness. This is consistent with the observations that metals explorations in 1986 was characterized not by announcements of "grassroots" discoveries, but by on-property exploration on identified mines or prospects. In an effort to cut costs, gold mining companies have been following up on low cost production opportunities in their prospect inventories. Short-term, low risk, heap leach operations that can produce rapid returns on investment have been pursued. Specific exploration projects are covered in the individual state sections of this Annual Review. Even though economic growth continued steadily through 1986, the domestic mining industry has

Jan 5, 1987

Power cable costs are only a small part of total mining costs. So many mine operators consider power cable failure and resultant downtime as part of the cost of doing business. But, viewed in terms of lost production, these costs can be quite significant. Now one company, Cablec, seeks to cut cable costs by upgrading the polymer compounding process used to make cable insulating and semiconducting materials. Cablec is the leading manufacturer of electrical power cables in North America. And with about a third of the market, Cablec is the largest supplier of power cable to the mining industry in the United States. To improve its products, Cable has entered the polymer compounding business. In July, it began producing insulator and semiconductor polymer compounds at its plant in Indianapolis, IN. "This new facility provides a quantum leap over conventional compounding methods," said Harry C. Schell, Cablec's president and chief executive officer. "The Cablec polymers plant is producing a dramatically higher standard of polymer compounds that provide significantly higher levels of performance and improved life cycle costs for power cable." Cablec faces tough foreign competition in the wire and cable business. Competing on price alone is difficult, particularly when foreign producers are state subsidized. So Cablec feels the best way to compete is to establish new quality production standards. The company's new polymers plant is one way to do this. By increasing purity control and uniformity in polymer compounding, Cablec says its power cables will last longer and fail less often. A typical medium voltage cable consists of a conductor, conductor shield, insulation, insulation shield, metal shield, and jacket. The conductor shield and the insulation shield are conducting polymers. Contaminants and imperfections can occur within the insulation, at the conductor shield/insulation interface, or at the insulation shield/ insulation interface. Over time, these contaminants and imperfections can decrease the electrical strength of the cable or cause premature cable failure. The effort to minimize the number and size of any possible contaminants begins with pure polymer compounds mixed in a clean facility. However, most power cable manufacturers manually handle raw materials, use ethylene/propylene (EP) in bulk bales, and mix polymercompounds in open Banbury mixers. The quality and uniformity of polymer compounds is also impacted by temperature variations in the mixing process. This results in wide gradations of product consistency from batch to batch and ultimately contributes to power cable failure. Cablec says the improved polymer compounds from its state-of-the-art plant will be the purest and most consistent insulating and semiconducting materials available. The plant itself RCA spent $18 million to build Cablec's Indianapolis plant. RCA used the facility to mix specialty polymer compounds used to make video disks. RCA had two considerations in mind for the plant, cleanliness and uniformity of the compounds. However, when the video disk market failed to materialize, RCA sold the 46.5 dam 2 (50,000 sq ft) plant to Cablec for $3.1 million. Cablec invested an additional $3 million for modifications and increased production capabilities. Today's replacement cost for such a facility is estimated at $30 million. Cablec says the plant will set a new standard for performance and be economically difficult to duplicate anywhere. One of the essential elements of the plant's clean process environment is the air intake system. It filters contaminants greater than 2 um, less than one-fiftieth the current industry standard. All material handling and conveying areas in the facility are air-locked. This keeps out contaminants such as smoke, dust, and pollen. Banks of pneumatic pumps move polymer components through the system and continually filter the air. The plant also has a backup air intake system. No process downtime due to pump failure here. From the time raw material enters the plant, it is stored, transported, and processed in filtered air by an airtight stainless steel system. The stainless steel resists rust and corrosion. This further eliminates the danger of contamination from paint or rust particles in the conveyance network. A computer system allows a single operator in a central control room to monitor every aspect of the compounding process from air quality to line speed. The computer

Jan 12, 1988

By J. M. W. Mackenzie

Three major developments have dominated the recovery of gold from ores, viz; 1846 Elsner's studies of gold solubility in aerated alkaline cyanide solutions 1887 MacArthur and Forrest's process based on zinc precipitation 1980+ Large scale development of the CIP process The brevity of this list for a metal whose extractive metallurgy has been as widely studied as any metal reflects the fact that the chemistry of the gold recovery process, as opposed to the engineering has seen few significant breakthroughs in over a century of application. to The purpose of this paper is introduce a new chemistry for gold recovery which may, in the future, add significantly to the technology of gold recovery. This new chemistry is based guanidine functionality - (R2R2N) 2 C = N R2 on the In developing the use of the guanidine functionality for the recovery of gold as the aurocyanide anion from leach solutions and pulps, Henkel has, as its objective, the development of an effective gold selective ion exchange resin. As will be seen, the initial investigation of the properties of the guanidine functionality in the solvent form for gold recovery lead to some significant metallurgical developments of their own but, the foremost objective was, and remains, the development of a resin for gold recovery. As a part of this development project, Henkel has entered into association with Mintek, South Africa, and this ongoing association has the potential to develop RIP/RIS technology for the gold industry. More details on Henkel's gold recovery technology and Mintek's role in this programme are given in (1). an The application of ion exchange resins to gold recovery is not new and many reviews have been published on the advantages of this process, and on the current state of the art of gold recovery by ion exchange resins. (2)(3)(4) Widespread application of ion exchange resin based processes for gold recovery has, however, not occurred to date. Fleming and Cromberge (5), described some significant pilot plant RIP trials in South Africa and the partial success of these pilot plant trials held out the tantalising possibility of a breakthrough in gold recovery based on RIP technology. Despite this pioneering work, only two really significant applications of RIP technology for gold recovery exist today. These are the Golden Jubilee Plant in the Eastern Transvaal South Africa and the widespread, although in some respects technically backward, use of ion exchange for gold recovery in the former USSR. There are two main reasons for the current scarcity of RIP and RIS plants used for gold recovery. These are: 1. 2. The perceived problems of engineering RIP plants The lack of a resin having suitable chemistry for both gold loading and gold elution The problems of engineering RIP plants have been largely overcome in the last 25 years. Much of the development in RIP plant design took place in the uranium industry during the rapid expansion of that industry in the mid 1970's. In addition, the development of CIP engineering systems has lead to the development of in-pulp solid extractant based metallurgical plant designs which can, with limited modification, be adopted to RIP. The lack of a resin having suitable chemical properties has almost certainly been the main reason for RIP/RIS not being introduced into the gold industry over the past decade. Attempts to use RIP/RIS have focussed on either conventional we ak base or conventional strong base resins. The weak. base resins have the disadvantage that they

Jan 1, 1993

By K. -A. Björkstedt

INTRODUCTION Strassa lies in the central part of Bergslagen, a tradi¬tional mining district, on the eastern side of the Stora Valley at an elevation of about 200 m above sea level. A railway siding runs between the mine and the Stora railway station from which there are railway connections to the shipping port and iron and steel works in Oxelosund, about 224 km away. The distance to the provin¬cial capital Orebro is about 60 km. The climate is typi¬cal for this part of central Sweden and is illustrated by the diagram of monthly precipitation and temperatures for the years 1968-1975 (Fig. 1). HISTORY There is no certain information as to when the Strassa mine was first worked, but it is known from sur¬viving accounts of mine inspectors that there were smelt¬ing works in operation in nearby villages in the 12th century. An example is the Gusselhytta ore smelting works, 10 km south of Strassa, which dates from this period. Around the year 1540 there were two smelting works in Strassa, the Upper Karberg and Lower Karberg works. Ore for these smelters was probably taken from Strassa and from the adjacent Blanka mine. In the year 1624 Strassa is mentioned by the painter Jons Nils Krook in an account of the iron mines in the Linde mining district (Linde Bergslags Jarngruvor). Several mines were listed in the area, the deepest being about 30 m. An impressive power installation is mentioned in 1639, including a piston system of lashed poles for transmit¬ting power from the Stora River to the Strassa fields. Its length was 2670 m. Common ground comprising about 20.2 km2 (5000 acres) of forest was allocated in 1689 for the furtherance of mining operations. Until the beginning of this century only the rich cen¬tral parts of the ore body were mined and these yielded, after handpicking, lump ore suitable for smelter feed. An example of the ore grades from these early times is an analysis of ore from the "Big Mine" (Storgruvan) from the year 1873: 48.5% Fe, 0.008% P, and 0.06% S. This same year a total of about 18 000 t was ex¬tracted from the Strassa mine. OWNERSHIP The mine was owned and run until 1874 by a min¬ing association made up of 119 so-called "bergsman," who were homesteaders often engaged in agriculture and timber-cutting as well. In that year the Strossa Grufvebolag (Mining Co.) was founded. In 1906 it was con¬verted into a joint stock company, the Strossa Gruveaktiebolag. This was acquired in 1907 by Metallurgiska AB for the implementation of Gustav Grondal's beneficiating and briquetting methods, for which the Strassa ore was well suited. The same year saw the completion of a new ore dressing plant with an annual production of 46 000 t of ore concentrate. In 1911 the mine passed to new hands, and in 1913 it was purchased by an Austrian company. Extensive new installations were made and in 1915 a new dressing and briquetting plant was completed with twice the capacity of the old one. In 1917 the Strassa mine was acquired by Granges. Be¬cause of unfavorable business trends and technical diffi¬culties, mining operations were brought to a close in 1923. Pumping kept the mine free of water until 1933 but it was completely filled ten years later. Up to 1950 the surface buildings and installations remained intact but the large dressing and briquetting plant burned to the ground in that year. Today only the machine shop re¬mains from this earlier period of operation, now housing parts of the Mineral Processing Laboratory. The decision to take up mining operations again was made in 1955 and construction work began the follow¬ing year. Of the old installation, only the "southern shaft" could be used for some development drifting after it had been completed with a new headframe. Other¬wise, all the buildings and installations required for the operations had to be rebuilt. New installations ready by 1960 were office and personnel facilities, a new shaft and headframe, a sorting and concentrating plant, a macadam plant, settling basins, pump stations, and a railway and yard with transport equipment. The instal¬lation was completed with two plants

Jan 1, 1982

By D. A. Norrgran, J. A. Marin

Introduction The recent development of rare earth permanent magnets has revolutionized the field of magnetic separation. The advent of rare earth permanent magnets in the 1980s provided a magnetic product that was an order of magnitude stronger than that conventional ferrite magnets. This allowed for the design of high-intensity magnetic circuits that operated energy free and that surpassed electromagnets in the strength and effectiveness. New applications and design concepts that focused on the mineral and metal processing industries have evolved. This technology led to the development of various magnetic separators specifically designed for mineral processing applications. Applications that were not previously considered are now being used in primary mineral upgrading, recycling and secondary recovery. Historical perspective Lodestone was the first naturally occurring permanent magnetic material known. Lodstone was most likely used to upgrade iron ore by early civilizations. By the 1600s, the early magnet technology had advanced to quench-hardened iron-carbon alloys. The practical significance of magnetic separation was formally recognized in 1792 when an English patent was issued for separating iron ore by magnetic attraction. By today's standards, carbon steel is a very poor magnet material. It is easily demagnetized and has a very low energy product of much less than I MGOe (Million-Gauss-Oerstads). This was state-of-the-art technology for almost 300 years until chromium was added to magnet feedstock, which resulted in a three-fold increase in the energy product. The well documented addition of cobalt to permanent magnets in 1917 initiated the 30-year era of "Alnico" magnets that at the time provided a superior magnetic energy product. Since then, the science of magnetism has advanced rapidly and is now considered a highly developed branch of physics and material science. Permanent magnets have had an extremely long history. Figure 1 presents a chronology of permanent magnets that illustrates the increase in energy product. Amazing developments in material science have taken place in the last two decades. The gradual advancement of permanent magnet technology was shattered in 1967 with the initial development of samarium-cobalt (rare earth) magnets. Since that time, the advent of neodymium-boron-iron magnets provided such an increase in energy product that new design concepts were considered. New avenues of study were introduced by the complexities in the material science and physics involved in describing these new permanent magnets. Furthermore, applications for permanent magnets that were previously not considered were now viable. Rare earth elements Rare earth elements have claimed the attention of scientists for the past century. These elements were originally termed "rare" because they were thought to be quite scarce. Since then, however, geological studies have shown them to be relatively abundant. The discovery and identification of rare earth elements is complicated by the inherent difficulties in separating them from each other. The rare earth elements comprise the fifteen transition elements of Group IIIB, Period 6, of the periodic table. These elements extend from lanthanum to lutetium and are commonly called the lanthanide series. Samarium and neodymium are the two most common elements used in the commercial manufacture of rare earth permanent magnets. Commercial grade rare earth magnets There are only a few common types of rare earth magnets that are considered in the circuit design for magnetic separators. Early rare earth magnets of commercial significance (introduced in 1970) consisted of the first generation of sintered SmCo5. The energy product of these magnets ranged up to 23 MGOe, which provided the initial impetus to the field of high-energy permanent magnets. Although these magnets did not produce the extremely high magnetic field strengths of current rare earth magnets, they were relatively temperature stable. Containing 66% Co, they are the most expensive of the basic commercial rare earth permanent magnets. Their use is limited today because they are being replaced by second and third generation rare earth permanent magnets.

Jan 1, 1995

By R. Kent Comann

Introduction This article describes ways to maximize employment potential for those looking for a job in the minerals industry. Advice on the right way to prepare a resume is followed by tips on where and how to look for a job and ways to make the best of an interview. Information on current salaries is also included. RESUME NAME: Give home and college addresses and telephone numbers, ADDRESS: if still in school. Otherwise, give home address and TELEPHONE: telephone plus business telephone number. JOB OBJECTIVE: What do you want to do? EDUCATION: List degree(s), name of college(s), location(s), and date(s) degree(s) received or to be awarded. Do not list your high school but do include postgraduate education, training, or seminars. WORK EXPERIENCE: List your most recent experience first, and the rest in reverse chronological order. Include dates (month and year) and company names. First show the entire time spent with each employer, then list the various jobs (with dates and duties) you may have had with this employer. MILITARY SERVICE: List dates, branch of service, and duties. If you did not serve, delete this section. COLLEGE ACTIVITIES: Show organizations, sports, and offices held. Only list these if you are a graduating senior, otherwise delete this section. PERSONAL DATA: List age, marital status, height, weight (optional), foreign languages spoken, professional memberships, and health (if good or excellent, otherwise do not mention it in the resume). If you prefer a certain location and would not accept a job anywhere else, list the location desired. Otherwise, do not restrict yourself. Also list special skills, hobbies, or interests that will help sell you to the company. Be brief in this section and use good judgement as to what you say. REFERENCES: Furnished on request. Resume Guidelines There are some things to remember when preparing a resume. These include: • Limit your resume to one page, unless you have a lot of experience. Then it should be a maximum of two pages. • Use 8.5 x 11 plain bond paper, white or off-white. Stay away from colored paper, odd size paper, and cute resumes. They are counter-productive. • Make certain your resume is typed professionally. Be sure to check for spelling, grammar, and typographical errors. • Have your resume reproduced by copier or printing shop. Do not use carbon copies. • Keep your sentences and remarks short and punchy. Use action verbs. • Have a good layout that is easy to read, brief, and uncluttered. Allow some white space in your resume. There are also some things you should not do when preparing a resume. These include the following: • Do not include a photograph or mention your religion, race, or sex. • Do not leave gaps of time unaccounted for in your resume. • Do not include your college transcripts with your resume or list all the courses you took in college. • Do not list all the articles you have written. Only list one or two, if appropriate for, say, a research and development job. • Do not include reference letters with your resume. Any good, prospective employer will make telephone reference checks. • Do not use "Curriculum Vitae," use "Resume." • Do not include unnecessary data such as your social security number, passport number, or PE registration number. If you are a PE or EIT, be sure to mention that in your resume. Do not mention US citizenship or US work permit unless you have a foreign education or are not a native US citizen. • Do not include present salary or salary requirements. This will come out in the interview. • Do not use the word "I" in your resume or refer to yourself in the third person. • Do not lie, distort, exaggerate, or editorialize in your resume. • Do not have your resume prepared by an outside resume service. This is unnecessary, costly, and generally not effective.

Jan 2, 1985

By Duane A. Smith

Horace Tabor. No 19th century Colorado mining man is better known but, unfortunately, probably less understood. He is little appreciated for his significant contributions to the industry and the state. A century ago, Tabor was headline news. The legend today, however, has become so all pervasive and so interwoven with fact that the two are hard to separate. For example, he never told Baby Doe on his death bed to hang on to the Matchless mine. Yet this has emerged as staple Tabor fare. So much attention has been focused on his matrimonial triangle that it overshadows the man. Symbol of mining's reward Tabor traveled west in the 1859 Pike's Peak gold rush. In the next generation, he came to symbolize the rewards mining might lavish on an individual. The Leadville Daily Herald (Sept. 10, 1882) could write, without exaggeration, that "the extent which the mining industry of Colorado is under obligation to Tabor cannot be easily estimated - what he has done for Leadville and Denver is patent to all." It had not come easily, nor had Tabor started out a success. His early placer mining at Payne's Bar, near present-day Idaho Springs, CO, had turned no fortune. So in 1860, he, his wife Augusta, and son went south to Colorado City, then over Ute Pass and up the Arkansas Valley to Oro City. Arriving soon after the initial discovery, Tabor staked a good claim. With a sharp eye, he and Augusta broadened their base by establishing a store. Both mining and business would pay dividends in the following years. By the season's end, though, Oro City was already declining. Always on the lookout for a richer district, Tabor and his family moved the next year across the mountains to promising Buckskin Joe. The familiar pattern followed, with the store and post office being the center of their attention. Mining investment and management now replaced the earlier physical panning and sluice operating. Seven years later, the Tabors abandoned Buckskin Joe and returned to Oro City. It had moved up California Gulch to be near the district's best mine, the Printer Boy. Middle class was not enough Middle class respectability, plus a steady income, was the Tabor's by the 1870s; fine as far as it went. But it proved far from the fortune Horace had always been seeking. Oro City languished in the backwash of Colorado mining and Tabor seemed like many men who had drifted around the Territory following the ebb and flow of mining. His faithful wife Augusta had been a steady factor in the success the family achieved. She helped year after year to operate the store and post office. Not simply the happy-go-lucky individual he has often been portrayed, Tabor was a hard working businessman/mine owner. An R. G. Dun and Company agent evaluated him in 1876 "Net worth $23,200. Is a very shrewd businessman and not liable to lose money, has a good chance to make money as he had no competition." Leadville: Tabor's silverlined-fortune After all those years on the Colorado mining frontier, in 1877, Lake County's wealthiest and most respected merchant made another move. It was short in distance, only 4 km (2.5 miles) down California Gulch, then a little north to a new mining camp. This new camp, soon named Leadville, gave birth to the Tabor fortune and legend. Middle-aged (Horace was 47), the Tabors once more set up their general store and found themselves in the midst of the open¬ing of a new district. This time silver beckoned and not the gold that had brought them West 18 years before.

Jan 1, 1989

By Gregory R. Wessel

Sulfur is a nonmetallic element widely distributed in nature and of great physical and economic importance. It is the 14th most abundant element in the Earth's crust (0.06%) and an important constituent of animal and plant life. Sulfur has been known and used since ancient times for a number of medicinal and industrial uses. At present, most sulfur is used to generate sulfuric acid that is used in a wide variety of industrial processes, particularly the production of fertilizer. Because of this, sulfuric acid (and hence sulfur) consumption is often regarded as a good index of a nation's industrial development. In the past, sulfur was mined from surface occurrences in several geologic environments, and was used in relatively small amounts. With time, the uses of sulfur and sulfuric acid expanded, as has the need for larger quantities of these commodities. Sulfur is now mined from both surface and underground deposits, and is recovered as a byproduct from a number of industrial processes. Despite valiant efforts and years of work by sulfur explorationists and others, many aspects of sulfur mineralization remain controversial. Almost as controversial is the spelling of the word sulfur. The English spelling sulphur commonly is used outside America and in the American sulfur mining industry, but sulfur is the correct American spelling as approved by the American Chemical Society, the American Geological Institute, and many others. Those new to the American sulfur industry often find it puzzling to be reprimanded for using the correct American spelling. Sulfur resources are abundant and widespread, but the extent to which they can be classified as reserves is constrained by pre- vailing prices and extraction technologies. At present, sulfur can be economically mined from very few deposits. The sulfur industry is roughly divisible into two sectors: voluntary (or discretionary) and involuntary (or nondiscretionary). In voluntary production, the mining of sulfur or pyrites is the sole objective, and the recovery of the resource is as complete as economic conditions will allow. During involuntary production, sulfur or sulfuric acid (termed recovered sulfur) are produced as byproducts, and the quantity of the output is dictated by the demand for the primary product. Voluntary sulfur now accounts for only about 35% of the elemental sulfur produced worldwide, and most inves- tigators believe that voluntary sulfur will be less important in the future. Sulfur sources and products are described as follows (after Barker, 1983): Sulfur Sources: Combined sulfur-sulfur that occurs in nature combined with other elements, commonly referring to sulfides and sulfates. Cupriferous pyrites-pyrite containing minor amounts of cop- per sulfides. Hydrogen sulfide-a toxic gas that occurs in petroleum and natural gas. Involuntary sulfur-sulfur produced as a byproduct in response to legislative or process mandates. Native sulfur-naturally occurring elemental sulfur. Nonferrous metal sulfides-opper, lead, zinc, nickel, and molybdenum sulfides that are processed for their metal content. Organic sulfur complex organic sulfur compounds that occur in petroleum, coal, oil shale, and tar sands. Pyrites-iron sulfide minerals that include pyrite, marcasite, and pyrrhotite. Sulfate sulfur-sulfur contained in anhydrite and gypsum. Voluntary sulfur-sulfur produced in response to market demand. Basic Sulfur Products: Acid sludge-contaminated sulfuric acid usually returned to acid plants for reconstitution. Brimstone-synonymous with crude sulfur. Bright sulfur-crude sulfur free of discoloring impurities and bright yellow in color. Broken sulfur-solid crude sulfur crushed to -8 in. Byproduct sulfuric acid-sulfuric acid produced as a byproduct of a metallurgical or industrial process, generally relating to combined sulfur sources. Crude sulfur-commercial nomenclature for elemental sulfur. Dark sulfur-crude sulfur discolored by minor quantities of hydrocarbons, ranging up to 0.3% carbon content. Elemental sulfur-processed sulfur in the elemental form produced from native sulfur or combined sulfur sources, generally with a minimum sulfur content of 99.5%. Formed sulfur-elemental sulfur cast or pressed into particular shapes to enhance handling and to suppress dust generation and moisture retention. Frasch sulfur-elemental sulfur produced from native sulfur sources by the Frasch mining process. Liquid sulfur-synonymous with molten sulfur. Liquid sulfur dioxide-purified sulfur dioxide compressed to the liquid phase. Molten sulfur-crude sulfur in the molten state. Prilled sulfur-solid crude sulfur in the form of pellets produced by cooling molten sulfur in air or water. Recovered sulfur-elemental sulfur produced from combined sulfur sources (including byproduct hydrogen sulfide, but sometimes referring only to sulfur from fossil fuels) by any method. Slated sulfur-solid crude sulfur in the form of slate-like lumps, produced by allowing molten sulfur to solidify on a moving belt. Specialty sulfur-prepared or refined grades of elemental sulfur that include amorphous, colloidal, flowers, precipitated, wettable, flour, and paste sulfur. Sulfur ore-unprocessed ore containing native sulfur.

Jan 1, 1994

By Ivor Hawkes

INTRODUCTION High energy breaking is an alternative to using ex¬plosives in underground secondary breaking operations. It also is a means of upgrading conventional hand-held breakers, manual sledge-hammer breaking, and scaling bar operations. Major areas of application are in sec¬ondary breaking over grizzlies and at drawpoints. Other applications include breaking down ripping lips in longwall seam mining, scaling in stopes and rooms, general demolition work, and roadway maintenance. There is considerable interest in high-energy impact breakers for use in primary ore breaking, but, as of 1977, all such applications have been only experimental (duToit, 1973; Joughin, 1976; Wayment and Grantmyre, 1976). EQUIPMENT Essentially, a high-energy impact hammer is a boom¬mounted pneumatically or hydraulically actuated breaker. The machine basically consists of a piston that oscillates in a housing and impacts the end of a tool or moil thrust against the rock. The force applied to the rock primarily depends upon the impact energy of the piston-the higher the impact or blow energy, the greater the force and, thus, the greater the rock break¬age. Among drill and breaker designers, a common expression for blow energy is "force of blow." Hand-held breakers are limited to blow energies of about 140 J (100 ft-lb), because the operator is unable to handle heavier machines efficiently or to absorb the recoil energy resulting from higher blow energies. How¬ever, these restrictions do not apply to boom mounted breakers; machines with blow energies on the order of 4000 J (3000 ft-lb) and higher are available commer¬cially for underground use. There is considerable evi¬dence to show that increasing the blow energy also in¬creases the efficiency of the breaking operation; i.e., more rock is broken per unit of energy expended (Grantmyre and Hawkes, 1975). Thus, there is a trend to higher blow-energy machines, particularly where high¬strength rocks are to be broken. In relation to rock breaking, the blow rate of boom¬mounted impact breakers is not as important as it is for rock drills. This is because the breaker must be moved over the work surface between blows. The blow rate is governed eventually by the power supply, and typical blow rates range between 200 and 600 blows per minute. As a general rule, light blow-energy machines have higher blow rates than heavier machines. Table 1 lists most of the boom-mounted impact breakers that were available commercially during 1977, and it gives details of the blow energies and machine weights. Restrictions are placed on the blow energy by the machine weight and size, and by the strength of the boom. Typically, boom-mounted impact hammers have a blow-energy to mass ratio of about 1.5, with lower values for lighter machines and higher values for heavier machines. In addition to supporting the hammer weight, the boom also has to absorb the recoil energy of the blow, which can be on the order of 1400 J (1000 ft-lb) for large hammers operating in a horizontal mode. Interesting exceptions to the general run of impactors are the Joy HEFTI hydraulic hammers. In these machines, the piston impacts onto a fluid cushion that is positioned between the piston and the impact tool. This approach allows very high piston velocities, over 30 m/s (100 fps), to be used without the risk of break¬ing the piston or impact tool. Steel on steel impacts must be limited to impact velocities of about 10 m/s (35 fps) due to the high impact stresses generated; thus, increased blow energies can be achieved only by increas¬ing the piston size. The Joy 514 HEFTI®, listed in Table 1, has a blow energy of 27 100 J (20,000 ft-lb), but, as of 1977, the machine has been used underground only on an experimental basis. Using a fluid cushion between the piston and the impact tool allows the use of light pistons, reducing the overall machine weight. The recoil energy, which must be absorbed by the boom for a given blow energy, is directly proportional to the piston to machine mass ratio, and operating with light pistons provides an addi¬tional benefit in reducing the requisite boom size. Both pneumatic and hydraulic hammers are avail¬able commercially. Although hydraulic hammers are a relatively recent development, they already outnumber the pneumatic machines in use. There are many reasons

Jan 1, 1982

By M. Kalin, C. Steinberg

We have worked on several flooded pits from coal-mining activities in the former East Germany, as well as ones associated with hard- rock mining, including the B-zone pit discussed in the above technical paper. We found the paper to be a useful summary, but, unfortunately, it failed to give an adequate comparison of the physical limnology of the flooded pits, which is an essential component. While the title suggests that the primary focus of the review is physical limnology, it appears that it is essentially pit-lake chemistry being presented. Physical limnology requires that factors such as fetch, latitude, light penetration, relation to ground water table, methods of flooding and the physical shape of the pits be defined. These physical aspects of a pit interact with the chemical and biological processes taking place in it, all of which contribute to the character of a water body. Few of these physical aspects are presented, however. The conclusion that the authors reach suggests that meromixis may be a condition that would serve as an effective containment mechanism for contaminants in a pit. Although this may be desirable, such limnological conditions are not clearly supported by the data presented for any of the pits. These data should be summarized to facilitate comparison between the same structural units of the pit water - the epi- and metalimnion for example. The thermocline depth is a reflection of the physical forces mixing the water body, and pit dimensions affect these forces. Due to the use of different scales in Figs. 2 through 5, it is difficult to determine whether the thermocline is at the expected depth, because the fetch is not given. Moreover, the status of a water body cannot be determined unless measurements cover a period of at least one year, and depth profiles are completed to represent the entire depth of the pit. This shortcoming is most notable in the case of the Berkeley pit, where data are given for depths of only 20 and 35 m (66 and 115 ft), although the pit is reported to be 242 m (794 ft) deep. Limnological data to define the status of the pit water have to be collected at regular intervals, for the same parameters. The authors present temperature measurements for 1-m (3.3-ft) intervals, but fail to use that interval for other parameters, such as dissolved oxygen or, in some cases, for contaminant concentrations. Furthermore, the profiles for the deepest part of the pit display only part of the picture, because pits are rarely conical. Profiles can be considered to represent the status of a water body only after other stations in the pit have been monitored regularly and the consistency is determined. For example, fresh water, which can enter a pit at any depth, would interfere with the proposed meromictic conditions. Similarly, organic material at the bottom of a pit, such as the fish-waste deposited in the Gunnar pit, contribute to oxygen consumption. Oxygen depletion alone is not indicative of meromixis. It is interesting to note that the Dpit arsenic concentrations could possibly be slightly higher than the B-zone pit concentrations at depth, although this is difficult to determine accurately when a log scale is used for the D-pit and not for the B-zone pit. In our investigations, we noted arsenic removal in the B-zone pit bottom water, which was due to the formation of particles that are relegated to the newly forming sediment in the bottom of the pit. Particle-carrying contaminants form due to a combination of geochemical and biological factors and TSS contributed from erosion of the upper parts of the pit walls, whereas the settling out of particles from the water column is controlled by the physical conditions or turn over, for example. during ice cover in the B-zone pit. Although meromictic conditions for flooded pits may be desirable at decommissioning, this would depend largely on the physical conditions of the pit, because, under no circumstances, would this water be of desirable ground-water quality. Under meromictic conditions, acidity, if an environmental issue, may be reduced by microbial acid-neutralizing activity, and several heavy metals may form more or less stable sulphitic compounds. These may stay suspended in the water if conditions are such that they are not relegated to the sediments, i.e., in the absence of turnover. These processes do not take place in meromictic conditions only, but meromixis does require autochthonous and/or allochthonous organic substrate supplies, which are generated under aerobic conditions. Specific limnological (biological, chemical and physical) features of the pit lake under consideration have to be defined, such that water quality parameters can be predicted, and the objectives of the decommissioning activities, environ-

Jan 1, 1999

By Donald C. Moore

Introduction Asamera Minerals (US) Inc. and its joint venture partner, Breakwater Resources Ltd., discovered ore grade gold mineralization on their 20 km2 (5000 acre) Wenatchee, WA land position in February 1983. Due to the high grade nature of the discovery ore and the known reserves of ore in the "B Reef' and "B West" zones previously outlined by other companies, a decision was made to construct a mine/mill operation near the known ore occurrences. Further drilling in the discovery area quickly expanded known gold occurrences to more than 3.6 Mt (4 million st) with tentative in-place ore grade of 7 g/t (0.25 oz per st) and minor silver values. Based on existing knowledge of the ore body and the rapidly increasing ore reserve, a decision to build a 1.8-kt/d (2000-stpd) mine and mill complex was made in the second quarter of 1983. A schedule was devised to begin immediate mine development, shaft sinking, environmental and land use permitting, and mill and tailings dam construction (Fig. 1). Meeting the scheduled startup date, April 1, 1985, required a fast track schedule in all areas. To this end, Asamera purchased the Oracle Ridge Partners concentrator. This was an assemblage of new equipment designed for use as a copper concentrator in southern Arizona. The purchase contained all of the major mineral dressing equipment - crushers, screens, rod and ball mills, etc. and an engineering package. It did not include most of the other required items, such as buildings, conveyors, pipelines, tanks, and pumps. At the same time, core samples were sent to two independent process development laboratories for initial flowsheet development. Due to the refractory nature of the carbonaceous ore, cyanide leaching was not feasible. Flotation was selected as the concentration process. Further testing showed that autoclaving of the flotation concentrate followed by cyanidation would result in overall recovery of about 85% gold. A mine manager was hired to begin assembling an operations staff, hire an environmental consulting firm, and begin mine development. Environmental and land use concerns were major obstacles due to the mine's close proximity to a city of 20,000 people. These concerns had to be rapidly defined so as to mitigate any adverse impacts from and mining processing operations. Baseline data dealing with weather, air and water quality, and sound were measured before start of mine construction. Concentrator and flowsheet development remained static until October 1983 while definition drilling and mine development proceeded. In late October, a process engineer was hired to coordinate development of a process flowsheet, purchase the remainder of the concentrator equipment, prepare a concentrator construction contract, finalize concentrator detail engineering, and combine environmental and process requirements with a tailings dam design. Process development There were only 17 months remaining to mill start up from the hiring date of the process engineer. Therefore, the process flowsheet had to be finalized rapidly. To accomplish this, samples of drill core from the highest grade (and therefore potentially the most commercial) ore zones were sent to an outside metallurgical laboratory to confirm beneficiation tests on the flotation process. Test results again showed that flotation would provide about an 86% gold recovery. Therefore, all further testing was concentrated on flotation and autoclave/cyanidation of flotation concentrates. Focusing on a well known process such as flotation was important in accomplishing the rapid design and construction of the concentrator. If, during these next phases, we were continually changing design concepts, layout, and process flow, the mill startup would have been delayed many months. Once a process flowsheet is selected the process engineer must obtain the process criteria needed to design the beneficiation system. For example, it was known in early December that the Oracle Ridge rod and ball mills were too small to grind 1.8 kt/d (2000 stpd) of Wenatchee ore. A decision had to be made to purchase a large, used ball mill and convert the Oracle Ridge ball mill to a rod mill. The process engineer must be cognizant of the process criteria needed to size and select equipment. If not, the process engineer must use the professional services of the equipment manufacturing companies to review the requirements that the equipment is asked to perform. For the Wenatchee system, this resulted in the adaptation of a ball mill to a rod mill with a weight limit of grinding rods to protect the mill bearings and drive trains. When a decision is required, the process engineer has to present the facts and options in a manner that allows a rapid decision. This information must include costs, equipment availability, and effect on the construction schedule. At the Cannon mine, there were process development details that resulted in decisions similar to the ball mill purchase. These included an increased flotation residence time from eight to 25 minutes, an increased thickener area requirement, a high pressure tailings pumping system, and area constraints in plant layout. All of these decisions had to be timely and required assistance from manufacturers' service engineers, and knowledge of the alternate costs and effects on construction completion. Equipment procurement It was decided in early 1983 to build the ore milling facility with Oracle Ridge equipment, augmenting it with used equipment

Jan 2, 1989

By D. J. Brunner, S. Mathur, D. McKinney

With the advent of faster micro-processors, the use of numerical methods to simulate complex fluid dynamic phenomena in three dimensions for use in design has become prevalent in the automotive, and turbo-machinery industries. The Computational Fluid Dynamics (CFD) method divides the region of interest into small control volumes which form the mesh representing the physical characteristics of the problem, and uses the finite volume method to intergrate the equations for the conservation of mass, momentum, energy and species over each control volume. Recent developments in CFD software expedite mesh generation, and enable the use of unstructured grids, comprised of tetrahedral volumes in three dimensions and triangular areas in two. CFD more accurately represents complex geometries and allows for relative movement of meshes enabling simulation of multiple moving bodies. 'ibis paper presents two examples of how CFD simulation can be used to assess mine and tunnel ventilation problems formerly addressed by application of analytical solutions which were developed assuming ideal incompressible conditions. CFD simulation is used to evaluate the impact of varying the airflow in a descentionally ventilated airway on the layering along the roof of smoke and hot gases resulting from a vehicle fire. Control of the smoke layer is required to enable safe egress from the vehicle, particularly if the vehicle is for personnel transport, and to ensure control of the fire contaminants throughout the ventilation system. The airflow required to prevent layering against the ventilation direction, calculated from the Bakke and Leach relations (Bakke and Leach, 1962), is compared with the CFD simulation results. An evaluation of the pressures, generated as a vehicle enters a tunnel portal, using CFD simulation, is also presented for unflared and flared portal configurations. These simulation results are compared with predictions derived using an analytical method which assumes one-dimensional and incompressible flow. Results of the CFD simulation are presented in an animated video format. SIMULATION OF BACKLAYERING In designing a ventilation system for a transit tunnel, the ability of the ventilating air to control and prevent backlayering of smoke and hot gases resulting from a vehicle fire is of prime concern. The buoyant nature of hot smoke causes it to rise relative to the colder, fresh air provided by the ventilation system. If the vehicle fire occurs in a descentionally ventilated tunnel, the smoke may tend to move upgrade in a layer above the incoming ventilation airflow. The layer may become thick enough to engulf a substatntial part of the tunnel cross-section upgrade of the incident that comprises the evacuation route. This effect is termed "backlayering' and it is similar to the development of methane layers in mines for which most studies related to backlayering have been done. Prediction Techniques Analytical A number of studies have been conducted (Bakke and Leach, 1962) to define the characteristics of this phenomena and as a result have produced relations which are used both in the mine and transit ventilation fields to define the air velocities required to control layering. In the transit industry the air velocity required to prevent the backlayering phenomena from occuring during a vehicle fire is called the "critical velocity" (Associated Engineers, 1975) and is dependent upon a number of factors: tunnel height, cross-sectional area and grade; ambient air temperature and density; and the heat release rate of the fire. Common practice in transit ventilation design is to provide an airflow which meets or exceeds the critical velocity. In order to determine whether or not the critical velocity can be achieved with a particular ventilation system, a one-dimensional simulation of the tunnel network is typically performed using programs such as the Subway Environment Simulation program (SES) originally developed in the late 1970's (Associated Engineers, 1980). The results obtained from SES are compared to the critical velocity to determine the adequacy of the ventilation system. Computational Fluid Dynamics For the backlayering simulations, a commercial CFD code which has been used successfully in a wide variety of engineering applications, was used. It provides numerous options for modeling laminar and turbulent flows, multiple turbulence models, definition of multiple species and chemical reactions between them, a variety of boundary conditions (including constant pressure and constant velocity inlets) and the ability to apply user-defined FORTRAN subroutines. It includes the ability to model conductive, convective, and radiative heat transfer. FLUENT also permits the use of "body-fitted coordinates" to match the computational mesh or grid to complex real-world geometries. Computational Fluid Dynamics Model The model developed to simulate the backlayering phenomena is comprised of an airway of rectangular cross-section, 4 meters wide, 4.5 meters high, and 200 meters long. A laterally

Jan 1, 1995

By M. J. Howes, C. A. Nixon

INTRODUCTION A safe heat stress control strategy for an underground mine has three elements: Application of an environmental measure which reflects physiological strain with sufficient accuracy for the range of conditions encountered underground. Acceptance of a functional relationship between the environ- mental measure and human performance which is used to optimise the environmental conditions achievable with either ventilation or ventilation and refrigeration. A management control strategy based on the environmental measure which is designed to ensure that work in environments where excessive physiological strain may occur is prevented and corrective action is initiated. The environmental measure that reflects physiological strain is the link between the three elements and, since the turn of the century, the discussion of the merits of various indices has been prolific. One problem in selecting a suitable measure or index is the ease with which it can be physically obtained relative to accurately reflecting the physiological strain. For example, wet bulb temperature is simple to measure and, for a particular mining sys- tem, it may adequately represent physiological strain, however, it would not necessarily provide the same relatively safe measure in a different mining system. The acceptance of a measure which can be universally applied has been confounded by both development and predisposition. That is not to say that there is only one "correct" measure and all others are unsuitable. It is self evident that if the application of a particular index has resulted in adequate control, then that mea- sure is correct for that situation. However, an understanding of the limitations is necessary to ensure that adequate control is maintained as mining conditions change. Almost 100 years after the question of heat stress in mines started to be dealt with in a collective manner, an analysis of the available information is leading towards a general strategy to control this problem. In the paper, the developments in heat stress assessment are briefly examined and followed since the earliest published observations on the effect of heat in mines (Haldane, 1905), efforts to determine a relationship between an environmental measure and human performance are reviewed and summarised and the benefits of control strategies such as acclimatisation and shortened shifts are discussed as they relate to Mount Isa Mines. The results of testing the prototype air cooling power instrument are discussed and a heat stress control strategy outlined. HEAT STRESS AND AIR COOLING POWER The operation of the human engine is analogous to other engines where the conversion of chemical energy from the oxidation of fuel to useful mechanical energy is not 100% efficient. In a diesel engine it is about 33% and in a human engine less than 20% resulting in at least five times as much heat produced by the meta- bolic process as useful work done. Metabolic energy production is related to the rate at which oxygen is consumed and is about 340 W for each litre of oxygen per minute. Using measured oxygen consumption and an average body surface area of 2.0 m2, the approximate metabolic energy production associated with different mining tasks is (Morrison et al. 1968):- • Rest, 50 W/m2 • Light work, 75 to 125 W/m2 (machine, LHD or drill jumbo operators) • Medium work, 125 to 175 W/m2 (airleg drilling, light construction work) • Hard work, 175 to 275 W/m2 (barring down, building bulkheads and timbering) • Very hard work, over 275 W/m2 (shovelling rock) Heat balance is achieved when the rate of producing heat (the metabolic heat production rate) is equal to the rate at which the body can reject heat mainly through radiation, convection and evaporation. Heat exchange between the lungs and the air in- haled and exhaled is normally less than 5% of the total and there- fore usually ignored. Any heat not rejected to the surroundings will cause an increase in body core temperature. Since heat stress is related to the balance between the body and the surrounding thermal environment, the main parameters required to be known when determining acceptable conditions are those associated with the heat production and transfer mechanisms. These can be summarised as follows: Metabolic heat production rates (M - W) Skin surface area (A3) (and effects of clothing) Dry bulb temperature (t[ ]) Radiant temperature (t[ ]) Air velocity (V) Air pressure (P) Air vapour pressure (e [ ]) The rate of heat transfer to or from the environment depends on the equilibrium skin temperature t, and the sweat rate S,. These in turn depend on the response of the body to the imposed heat stress and the effect of thermoregulation (Stewart, 1981). Thermoregulation The body contains temperature sensitive structures which send impulses to the brain at a rate depending on the temperature. Both hot and cold signals can be differentiated and the thermoregulatory response ahivated according to which signal pre- dominates. If "cold" signals are dominant, body heat loss is re-

Jan 1, 1997

By James N. Grassby

Metal production has traditionally been a cyclical industry. 1977 in the base metals industry has highlighted that fact with a steady decline in the market place for base metals. Corporate aggregate planning is based on good forecasting and when changing conditions make that forecasting inaccurate, alternate analysis of schedules must be accomplished quickly. The computer is the only medium powerful enough to answer management's questions within an acceptable time frame. Worldwide communications networks now make it possible for the many multinational mining companies to arrive at coordinated corporate production plans based on similar basic data from several producing divisions. The Ontario Div. of Inco Metals Co., a unit of Inco Ltd., encompassed during 1977 14 producing mines, 2 nonproducing mines, 1 nonproducing open pit, 5 flotation mills, I smelter, 2 nickel refineries, 1 copper refinery, and an iron ore recovery plant. Several new mines are under development. Mine production approximates 63 500 tpd (70,000 stpd) from some 1000 workplaces. The need for longrange mine planning and scheduling is obvious and until recently has been satisfied by manual methods. However, while manual methods are often quite practical for a small number of mines, or for mines with relatively short lives, or for mines with only one mining method, the situation in the Ontario Div. is too complex and long-term to allow effective, timely planning by such methods. In 1967, a year after the first computer was installed in the division, work was begun on the computerization of the more than 30,000 borehole logs kept in the mines exploration files with a view to speeding up the assessment of mineral resources. Also in 1967 one of the first substantial systems developed on the main business systems computer was the monthly scheduling of development and production for all mines. This system encompassed tons productions per workplace, feet development per workplace, auxiliary and service activities and the requisite labor for each of these. These two dates are mentioned to show that concurrently with the acquisition of computers in the division there was an early understanding of the importance of automating mine scheduling and planning and of the potential for effective and useful computerization of such activities. In 1970, a pilot program was developed in one of the mines containing a large number of workplaces (350) to schedule the production on a network basis. Varying degrees of interest were shown by the individual mine engineers based on the complexity of the production planning problem at each mine. In an evolutionary process, mines were added to the system as its value became obvious in terms of time saving, flexibility, and accuracy. In 1975, the central mine engineering department decided that the system should be extended to all mines so that a complete division-wide long-range scheduling system would be available upon which to base development and production planning. This system would use basic ore data, allow the addition of scheduling commands, and calculate a schedule for the life of the mine. Many reports could be produced for each mine and all-mine summaries would also be available. The significance of the system can be stated in one phrase-speed, accuracy, and flexibility. Whereas in the past a long-range all mines schedule might have been done once a year, taking from two to three months, now it can be done in a day. The chief mines engineer can incorporate changes in input data and external constraints and call for a "new look" at the long-range schedule to assist him and mines production management in their decision-making processes. Using previous manual methods this could not be done quickly enough to be useful.

Jan 1, 1979

By Joseph R. Wojcik

Introduction Placer deposits account for one-half to two-thirds of the total gold produced worldwide. As many as 311 t (10 million oz) have been produced from some major districts. Several others are credited with 155 t (5 million oz). All the major districts and many lesser ones have similar characteristics. Consideration of these characteristics offers insight into the origin of gold in the deposit. Knowing the gold deposit's geologic history leads to useful conclusions in exploring for placer deposits and for deposits in consolidated areas. Placer Deposit Characteristics Table 1 lists 17 placer gold districts around the world and their characteristics. These include type and age of, mineralized source rocks, associated intrusive rocks, and age of intermediate sedimentary cycles. Although quartzite or sandstone occurs in the bedrock sequence in four of the 17 districts, slates, phyllites, or schists are present in all of the districts. Additionally, in all but three, the slates are pre-Mesozoic. And, in six of the remaining 14, the slates are Precambrian. Rocks intrusive into the slates are all of intermediate to acidic composition and are emplaced as stocks, dikes, or sills, rather than being batholithic in proportions. Intermediate interceptors have not been recognized in only three of the 17 districts. Data from China and Mongolia are inconclusive as to this characteristic. Black Shale Gold Higher than average gold values are found in modern sedimentary rocks in coarse granulites, conglomerates, sandstones, and gray-wackes. Not so predictably, higher than average gold values also occur in black shales and carbonaceous argillities. Li and Shokena (1974) found that, in the Proterozoic sedimentary and equivalent metamorphic rocks of the Enisei Ridge, carbonaceous and graphitic phyllites and schists were particularly enriched in gold. Korotayeva and Polikarpochkin (1969) reported gold content in organic rocks to be enriched up to 15 times more than the organic free shales and siltstones of the Nerchinsk Zavod region of eastern Transbaikal. Black shales can be gold interceptors as are rocks of the granulite facies. Gold occurs in shales in particulate, free form, in pyrite, adsorbed on clays, and precipitated on carbon. In copper-bearing shales of Poland's Fore-Sudetic monocline, Kucha (1973) reported gold present in organic compounds as well as in silver bearing minerals. Anoshin (1969) found a strong correlation between gold and the pelitic fraction of sediments in certain rocks of the Atlantic Ocean basin. Heavy metal absorption on clays results in agglomeration to the extent that the particles settle with the heavy mineral fraction in enriched strata. Piper and Graef (1974) think that gold in sediments on the flank of the East Pacific rise is of lithogenic origin. Crockett (1973), however, proposed a volcanogenic exhalative origin. A strong correlation between gold and volcanic activity exists in rocks of the Canadian Shield where tuffaceous sediments and porphyries are more frequently enriched in gold than rocks with no volcanic components. Volcanic rocks of intermediate to acidic composition are considered here to be the primary source of gold in black shales. Gold is concentrated in shales as detrital free grains with diameters less than 50 µm (270 mesh), as precipitates, and in pyrite or other sulfide minerals. This is the first step in the history of gold placer deposits. Grain Size Modification Gold, as free grains with diameters less than 50 µm (270 mesh), is too light to be concentrated hydraulically into placer deposits. Some process has increased the average size of the gold grains. Mobilization during regional metamorphism converts most of the gold into particulate grains or entrains it in microfractures in the sulfide minerals. With moderate folding, small, white, vitreous quartz veins develop. As deformation continues, larger, more conspicuous quartz veins are formed. The larger veins are folded and faulted and contain native gold as minute specks and flakes not visi-

Jan 11, 1984

By A. T. Yu

Two major events highlight recent developments in US ore and coal ports: completion of the last series of modern taconite pellet transshipment facilities on the Great Lakes; and modernization and construction of coal ports, particularly on the East and Gulf Coasts. The New Taconite Transportation System To reduce raw material transportation costs, a fleet of new generation high-capacity 304-m (1,000-ft) self-unloading vessels were built to carry iron ore pellets from the Minnesota-Michigan iron ranges to the steel plants on the lower Great Lakes. Existing port facilities had to be modernized, revised, or completely rebuilt to accommodate these large vessels. Some of these were Burlington Northern's Allouez, WI, loading dock; Duluth, Missabe & Iron Range Railway Co.'s Two Harbors, MN, facility; Republic Steel's Lorain, OH, facility; and Chessie's Toledo, OH, dock. Allouez and Two Harbors receive taconite pellets from unit trains and load them onto large vessels either after dumping or via a large stockpile and reclaim system. The Lorain facility receives iron ore pellets from self-unloading vessels' discharge boom conveyor and reloads the pellets into rail cars or small vessels destined to inland steel mills. The Toledo facility receives Armco pellets from vessels, stockpiles them through the winter, and reloads into unit trains destined for Armco's mills along Chessie's rail tracks. Burlington Northern's $75-million Allouez pellet dock, completed in June 1977, was built to receive pellets produced by Hibbing Taconite Co. and National Pellet Plant in Minnesota, stockpile them through the winter when the lakes are frozen, and load them into 304-m (1,000-ft) vessels in the shipping season. As much as 10 Mt (11 million st) of pellets may be stockpiled within the loop track. A 6-km (4-mile) long conveyor system connects the stockpile area and the dock. Thirty-six new concrete silos were built on the dock to house 2 kt (2,000 st) each of pellets before shiploading. The $35.5-million expansion of the Two Harbors transshipment facilities began shiploading in July 1978 after ground breaking in Aug. 1974. Particularly noteworthy is the first application of the Orboom system-a breakthrough in technology for the modernization of the century-old pocket docks on the Great Lakes to accommodate the new generation of super vessels. The pocket-type loading dock has been a standard on the lakes for nearly a hundred years. Bottom-dump rail cars fill the ore pockets on top of a finger pier. Gravity chutes matching standardized ore ship hatch spacings are lowered to fill the holds of a 20.3-kt (20,000-dwt) ship. The construction is simple and the loading swift. In spite of advances in technology elsewhere, most of these docks continue to serve the iron ore and coal trade in the same manner they did in the 19th century. Although performance of the pocket docks on small vessels remains outstanding, the new 304-m (1,000-ft) vessels are beyond the reach of the old docks. After extensive development, the Orboom system (Patent No. 4,065,002) for pocket docks was successfully developed. The heart of the Orboom system is the retractable shuttle loading arm which loads the wide beam vessels. The Orboom shuttles are fed by existing pockets of the dock that, in turn, are charged by a tripper conveyor along the length of the dock. The Two Harbors shiploading system is supported by a 0.9 Mt (1 million st) storage-reclaim network. Unit trains are bottom dumped. Taconite pellets are stacked and reclaimed by bucket wheel reclaiming systems. Lower Lakes New Ore Ports At the lower Great Lakes receiv-

Jan 10, 1982

By R. A. Metz, William H. Breeding

The remarks that follow are based substantially on experience covering 45 years, 80% of which has been in placer work, rather than on a review of available literature. Most commercial placers have been deposited by the action of water. The richer and more difficult-to-mine placers are those in the headwater areas where gradients are steepest. The most lucrative placers are generally in intermediate areas where volumes are greater, fewer boulders are present, and gradients are from 3% to 1-1/2%. The higher volume, lower grade placers are in the lower reaches of river systems where gradients are lower. Where gold-bearing rivers have discharged into the sea, wave action can concentrate values on beaches, past and present. Most of the rich, readily accessible placers were mined by our forefathers. Current opportunities exist: (1) in remote areas where infrastructure has been absent in the past, or development has been prohibited by adverse ownership - political or commercial; (2) in deposits that could not be mined by equipment available to our forefathers; (3) in deposits unidentified by our forefathers; (4) where the-price-of-product/cost ratio is substantially better than in earlier years; or (5) a combination of those factors. When I entered the placer business in the late 1930s, and subsequently, a prevailing opinion believed that glacial deposits should be avoided as irregular in mineral content and composition, and unrewarding to explore and develop; yet an operator has been mining a fluvio-glacial deposit profitably for the past 17 years. Rich buried placer channels, often called paleo-channels were worked in the last century, generally by hand methods, and under conditions that would be unacceptable today. Exploration and mining equipment now available make some of these channels attractive targets. Well-known examples are in California and Australia. The formation of a commercial placer requires a source of valuable minerals. Above primary deposits, there may be eluvial deposits formed by the erosion of gangue minerals and the concentration "in situ" of valuable minerals. Down slope from these deposits are the hillside or colluvial deposits, and below them are the alluvial deposits of redeposited material. Most of the great placer fields of the world are the result of several generations of erosion and deposition. Well-known examples are in California and Colombia. Gold is a very resistant and malleable material, and gold placers may extend for 64 or 80 km (40 or 50 miles) along a river system. Platinum is less malleable, but is very resistant to disintegration. Diamonds are extremely hard, and (especially gem diamonds) may be found over great lengths of a river system. Cassiterite is less resistant to disintegration, and tin placers seldom extend over two miles without resupply from an additional source or sources of mineralizaton. Tungsten minerals are generally more friable, and within a few hundred yards of the source disintegrate to the point that they are uneconomical to recover. Rutile, ilmenite and zircon placers generally result from the weathering of massive deposits, and may be encountered over extensive areas; most are fine grained and durable. What does a geologist or mining engineer look for in placer exploration? The old adage to look for a mine near an existing mine is still valid. You need a source of valuable mineral. Then you require conditions for concentration, which means a satisfactory gradient and/or other conditions that will permit heavy minerals to settle. Nicely riffled gravel, often called a shingling of the bars, is conducive to placer formation. Coarser gravel is logically associated with coarser gold. Excessive clay and/or high stream velocities in narrow channels can carry gold far downstream and distribute it uncommercially over a large area. When material is extremely fine, in situ weathering and concentration become more important. Placers frequently occur distant from lode mines, and one must remember that in a larger watershed the exceptional floods that occur once in a hundred or a thousand years can move great quantities of material long distances. The carrying power of water is said to vary with the fifth or sixth power of its velocity. I am not ready to disagree with Waldemar Lindgren and accept that many commercial placers are substantially enriched by the chemical deposition of gold from solutions; however, I have seen crystalline gold in clayey material quite distant from known sources of primary gold that is dif-

Jan 1, 1992

By William S. Gilmer

INTRODUCTION As man moved through the ages into the era of metals, he became increas¬ingly dependent on the extraction and beneficiation of various ore types. His progression has required more material, new sources, and inevitably, a need to extract the values from ores of much lower yield. Inherently welded to this need to extract has been the attendant requisite of power and water to carry out these tasks. With the richness of these deposits diminished, and the in¬crease of demand, mining has become so intimately tied with the needs of these auxiliary systems that the modern day milling operation can ill afford to ignore them. In the times of Agricola 16th century methods of milling required little thought for power or water consumption. Reduction was accomplished by hand¬sorting and hammer and grinding done with stamps and millstones. Medieval concentration was carried out by the "Jigging sieve" with agitation supplied by the sleeveless arm of the operator (The Mines Magazine,_ March 1933). Little consideration was likely given to the power source unless it became apparent that he had insufficient means to meet the demands of a dawn to dusk operation. Fuel for power production was also probably of little consequence unless sustained drought or warring neighbors burned the miners fields. Water was available from any adjacent stream with recycle, potable, tailings and mill consumption all going and coming from the same source. By the end of the 18th century and the early 19th century the requirements of larger throughputs and better re¬covery were forcing the miner to the need for greater sources of power and water. Arrastras, capable of accepting several hundred pounds of charge, con¬sisted of a clay bottom, a "liner of rocks harder than the substance to be crushed" and a revolving center post in which the ore was ground and amalga¬mated. Used in conjunction with a slucing system, this formed the primary recovery method of the time. Power was supplied by horse or oxen or, if terrain was amenable, by use of a horizontal or overshot water wheel. A typical system required 4.5 kw (6 hp) and needed to run in the vicinity of 15 rpm (Pre¬ston, 1895).

Jan 1, 1986

By Aladar B. Dory

INTRODUCTION Practical experience in mines with known presence of radon daughters in the mine atmosphere in Canada and elsewhere shows that a very high concentration builds up in an unventilated dead end heading. As Holaday et al1 observed, even a minimal air movement results in a drastic reduction in radon daughter concentration. It is therefore obvious that the main objective of radon daughter control in the working environment is to design the ventilation system providing an optimized flow of fresh air into the workplace, resulting in acceptable climatic conditions and achieving radon daughter concentrations resulting in exposures as low as reasonably achievable. BASIC OBJECTIVES Large mining companies, having extensive material resources and professional expertise, have utilized elaborate electrical modelling in the design of mine ventilation systems as early as 1950 (coal mining industry in Europe) and with the advance of computer modelling techniques, their utilization in ventilation systems design is on the increase. Unfortunately, these methods are usually not available to small mining companies and even the large companies might not achieve the fullest benefit from utilizing them, if proper limiting factors are not considered in the modelling. When an evaluation of a ventilation system of a mine is undertaken in literature, a measure of the amount of air supplied underground per one ton of ore mined is used as an indicator of the efficiency of the ventilation system. Yet, even the greatest amount of air forced into the mine might not result in an acceptable working environment if a proper distribution of this air into individual working places is not achieved. The volume and the age of the air are probably the two most important factors in achieving acceptable radon daughter concentrations in the workplace, but other factors also have to be considered. DIRECTOR MINE - ALCAN, NEWFOUNDLAND FLUORSPAR WORKS ST. LAWRENCE, NEWFOUNDLAND, CANADA Ventilation To illustrate the effects of the design of the ventilation system on the control of radon daughter concentration, let us review the gradual development of the ventilation system of this mine from the earlier years of its development up until its final years of operation. This mine, located near the community of St. Lawrence on the south coast of Burin Peninsula was developed in the late thirties and reached full production by 1942. Unfortunately as was customary at that time, the only source of ventilation was a natural draft. The mine was extremely wet, and no significant attention was initially given to possible health effects of dust. It was not until the mid-fifties, when a number of cases of silicosis had surfaced, that de Villiers and Windish2 observed a significant increase of lung cancer incidence among the miners in comparison to its incidence among the general population of Newfoundland. Suspicions regarding radiation as a cause of the lung cancer were expressed, but it was only in surveys taken in late 1959 and early 1960 that Windish3 and Little4 established the presence of radon daughters in the mine atmosphere in very high concentrations. Windish, de Villiers and Hurley suggested that the most likely source of the radon in the mine was the mine water which dissolved radon during its passage through the granitic country rock in the surrounding geological area. This conclusion was confirmed by analyses of water from various areas of the mine by the Atomic Energy Canada Limited laboratories. The radon values in the samples varied from 4,240 to 12,850 pCi/L5. Following the discovery of the presence of radon daughters in the mine, the company took speedy action to install mechanical ventilation for the mine. The system was not designed as a total unit, but fans were installed rather on a trial and error basis. The basic system installation began in March 1960 and was completed by 1962. It remained basically unchanged with only minor modifications until August 1973 when a wholly new, redesigned ventilation system was implemented. A schematic section of the mine and its ventilation system for the period prior to March 1960 is given in Figure "A", for the period 1960-1973 in Figure "B", and for the period after August 1973 in Figure "C". The ventilation system prior to 1960 is not known. All workings of the mine were ventilated only by natural ventilation. If any measurements of airflows at different or any times of the year ever existed, no records have been preserved. The very minimal natural ventilation was augmented by "blowing" air from compressed air supply lines and exhaust air from drills. It is known that the compressor capacities of the mine were limited and therefore no significant air movement was probably created by the "blowing".

Jan 1, 1981

By J. B. Gwiazda

Maintenance of a constant load setting throughout longwall support units and selection of the proper initial bearing capacity depending on the type of roof strata are the basic factors that ensure good performance of roof support in a longwall. These requirements can only be met by hydraulic support. The greatest advantage of hydraulic support is achieved when uniform pressure is imposed on the roof throughout the length of the longwall. Such support, however, is provided only if each of the support units acts with the same force against the roof, i.e., has the same setting load. In such cases, the roof behaves like a uniform plate without bending and shearing stresses, thereby ensuring an undisturbed structure. Without a positive set value system, achieving an equal setting load for all units of the longwall support is impossible. Due to alterations of the feed line pressure of support units as well as some reasons related to man's psychology, operators extend the height at different setting loads. This produces nonuniform roof stress, and disturbs the structure. Consequently, the roof usually cracks. The author has developed a positive set value system, which is described in this Technical Note. Selection of the setting load Two pressure values in the feed lines are usually applied in longwall hydraulic systems. Lightweight support is fed by 25 MPa (250 bar) liquid, while heavy duty units receive a nominal pressure of 31.5 MPa (315 bar). Such pressure is required not only for the props but also to power the adjust¬ment jacks and the advancing ram. If the feed pressure is too low, there will be difficulty in shifting the unit despite the inversion system of the advancing rams. On the other hand, for many roof types, the feed pressure often appears to be too high when applied as the setting load pressure. An excessive setting load acts too strongly against the roof, crushing weak strata close to the roof. The author has recognized a case where an excessive setting load destroyed not only the nearby roof strata but also the strata above a 2-m (6.6-ft) sandstone layer. In addition, an excessive setting load relieves the side¬walls, increasing the resistance when using cutting machines. As a result, the yield of coarse coal is diminished, and increased fines dominate in the final product, lowering its economic value. As indicated, selection of the proper setting load, depend¬ing on the mining and geological conditions of the extracted seam, is extremely important. In some mines, measures applied to prevent disturbances include reduction of the feed line pressure by adjusting the feed pump valve. The disadvantage accompanying reduced feed line pressure is more difficult operation in advancing the ram. Due to the reduced feed line pressure, the force of the advancing ram is much lower than the designed value. Other designs suggest using a third feed line. However, installation of supplementary valves on the support units is required, a time-consuming and expensive procedure. The disadvantages of the powered supports are eliminated by a system designed by the author. So far, such a method of setting load control has not been used in any type of support. Setting load control unit The designed positive valve set for prop loading and the setting load control correspond to existing control systems for hydraulic powered support. The layout of the unit connected to the hydraulic prop control is presented in Fig. 1. The unit is marked LIDS. It incorporates three valves that may operate separately or connected. Valve A automatically opens and closes with liquid flow in the prop feed circuit. The valve is opened when the canopy touches the roof and closed when the support unit is withdrawn. Valve B serves as the setting load control. Valve C automatically opens and closes the flow in the line connecting the under-piston space of E to the prop F with the separator G. The valve block of each support prop is marked BZ. The UDS unit is connected by the hydraulic lines to the F prop control circuit. A valve is connected by H to pressure line J and by K to the G separator. In the UDS-3 version, line L is connected to M, linking the over-piston space of F prop with the G separator. Valve B is linked with valve A by a connector; it is also connected to the under-piston space E of prop F by line P. Valve C is fixed between lines K and P, connecting space E of prop F with the separator G. When setting the support, the liquid flows from line J through separator G, the BZ valve, and valve C to the space E of prop F. When reaching the roof with the canopy, valve A is opened and C closes. In this way it is impossible for the operator to cause the liquid pressure in space E of prop F to reach the level of line J. The prop pressure is set by valve B of the UDS unit. When withdrawing the support, valve C is automatically opened and A closed. Three UDS units have been fabricated and are designated

Jan 1, 1990