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By J. W. Shield

Refractory sulfide flotation or gravity gold concentrates are coated onto a sized support matrix and placed in a heap configuration for biooxidation. The surface coating is typically less than 0.05 in. thick and the support matrix is preferably composed of 0.375 to 0.75 in. diameter material. The exposed thin coating of sulfides reduces biooxidation rate limitations due to oxygen mass transfer. The large spaces between matrix particles reduce resistance to fluid percolation and aeration in the heap. The capital and operating costs for this biooxidation configuration are significantly less than for concentrate biooxidation in stirred tanks. After 30 to 90 days of biooxidation, sub- sequent gold recoveries exceed those seen in whole ore heap biooxidation and are comparable to recoveries obtained from stirred tank biooxidation processes. Results from laboratory and pilot testing indicate that this technology is a viable alternative for pretreating sulfide refractory gold ores.

Jan 1, 1997

By N. Merriam

This report describes Koppelman's work to develop the K-Fuel® Series C Process used to dry and stabilize Powder River Basin coal. It also contains summaries of the results of bench-scale and pilot plant tests conducted to prove the Series C Process. The process is being used by the KFx-Thermo Ecotek partnership for a 500,000 ton per year commercial plant that has been constructed near Gillette, Wyoming. The plant is presently undergoing shakedown.

Jan 1, 1997

By V. O. Tenorio, J. Puell

"Dump design is critical to effective mine planning in large scale open pit mines. Capital and haulage costs are considerable from early stages and through the life-of-mine as these dumps gradually become immense structures. Delivered mining rates, as well as certain spatial and physical constraints, provide a set of parameters of mathematical and economic relationship that creates opportunities for modeling and thus facilitates the measuring and optimization of ultimate dump design by using programming and empirical techniques while achieving economic objectives. This paper presents a methodology to optimize the ultimate dump design by minimizing its total dumping costs. The methodology consists of the formulation of a generalized model based on the relationship of the main dump parameters, followed by the use of linear programming to find the solution that minimizes the haulage total cost, and finally through a series of iterations, modifies the preliminary solution by projecting it to the topography and creating the ultimate dump design. A case study applied to a waste dump exemplifies this methodology. INTRODUCTION Open pit mine operations move large amounts of material on a daily basis from the shovel faces to the appropriate destination points. Hustrulid et al. identified three major destination places where the material receives downstream treatment after its delivery: mill processing, leaching dumps, and waste rock dumps [1]. The corresponding destination is given by a cut-off grade criteria, ore type, and the possibility of obtaining the maximum metal recovery. Dump leaching technologies have developed substantially since 1975 [2], and have enabled mining facilities to beneficiate low grade sulfide and oxide ores. As a consequence, the number of active dumps have increased since then and the entire mine configuration, which is limited in itself, has changed. The proper location, design, and construction of a mine dump must serve the safety, environmental and economic goals of the long-range mine plans. Therefore, an optimized use of the available surface area in the mine is highly desired."

Jan 1, 2018

By Donald M. Ulbrich

Any of you who served in the military remember the admoniton to "get the lead out," In the military environment, where cost wasn't always a factor, "get the lead out" was more often a call for activity of any sort than activity of a particularly productive nature. As a result, when the top sergeant yelled "get the lead out," lots of people started moving.. . usually to get out of his way. Now there's a new cry, ..one that comes not only from the military but from an entire nation threatened with dependence on foreign sources for its energy needs. That cry ...get the coal out, Like the earlier order, it's a call to activity. A signal that the coal reserves of the United States must be tapped so that Project Independence can succeed by 1985, But it's more than a mere call to activity; it's also a call to achieve productivity. ,a call to produce coal both efficiently and acceptably. And that leads us to the key questions of where the coal is going to come from to meet that 1985 goal, and how can it best be mined. Where will the coal come from? There's no doubt much of it must come from surface mining in the West. How can this coal be mined? There are several options available but there is an excellent case to be made for using mining shovels and haul trucks due to their versatility in stripping, loading and reclamation, We'll explore this approach in this talk today.

Jan 1, 1976

By William G. Doepken

GENERAL DESCRIPTION The Henderson mine utilizes a highly mechanized, continuous panel-caving system to extract ore from a deep, massive, molybdenum deposit. The mine site, lo¬cated 80.5 km (50 miles) west of Denver, is situated at 3170 m (10,400 ft) above sea level in a steep valley on the eastern side of the Continental Divide. The mill site, located on the western side of the divide, is linked to the mine via a 15.5-km (9.6-mile) railroad tunnel. Mine production is scheduled to reach 27 215 t/ d (30,000 stpd) from the uppermost production level. This pro¬duction rate will have been reached within the four years following an intensive ten-year preproduction development program. A second production level, due to come on stream in later years, will expand production capabilities to 31 751 t/d (35,000 stpd). The mine, which employs approximately 2000, has blocked-out ore reserves of 220 Mt (242 million st) at an average grade of 0.42% MoS2. Geology The Henderson deposit is composed of two partially overlapping ore bodies that lie 1080 m (3600 ft) beneath the peak of Red Mountain. The ore bodies are entirely within a Tertiary rhyolite porphyry intrusive complex that has intruded Precambrian granite. The de¬posit is elliptical in plan with overall dimensions of 670 x 910 m (2200 x 3000 ft). In section it is arcuate with an average thickness of 185 m (600 ft). The top of the deposit is at an elevation of 2610 m (8700 ft) while the lower limits range from 2340 m (7800 ft) on the west to below 2100 m (7000 ft) on the east. The mineralization is relatively continuous in the ore bodies and consists of molybdenite and quartz in random to subrandom, intersecting, closely spaced veinlets. Cavability RQD (rock quality designation) * values measured from Henderson drill core and compared with the RQD values from the Climax and Urad block-cave mines in¬dicated that the Henderson ore would cave but with difficulty and that a high percentage of the cave ma¬terial would be blocky. Initial caving experience shows the Henderson ore caves easier and with greater frag¬mentation than was indicated by the RQD values. Pos¬sible reasons for this are: 1) The greater lithostatic load at the Henderson has a greater effect on caving and fragmentation than anticipated. 2) The relief of confining pressure due to caving action and brittleness of the rock is enhancing breakage. Mine Layout The Henderson mine layout was designed from its inception for a large-scale, rubber-tired, continuous panel-caving mining system. The ore body is vertically divided into three major mine levels (Fig. 1). The top¬most level (8100 level), which is 2469 m (8100 ft) above sea level, is the current mine production level. The intermediate or 7700 level, which is in the initial

Jan 1, 1982

By J. M. Van Langen

In 1965 Hoogovens was looking for additional capacity for burden preparation in the order of 2,5 to 3 million metric tons per annum of some kind of optimal blast furnace ore feed. In principle four solutions offered them-selves for further investigation: 1. purchase of screened ore and/or run of mine ores with the provision of a corresponding ore screening capacity on the blast furnace site; 2. manufacture of sinter; 3. purchase of pellets; 4. manufacture of pellets. After studying the technological, technical, economical and commercial factors Hoogovens decided for the last mentioned possibility. Before taking this decision not only a great number of different mixtures of iron ore were tested, but during a long period in one of the big blast furnaces a number of tests were also executed with purchased pellets from South and North America. It was found, that using 50% acid pellets with 40% basic sinter a coke rate saving of 25 kg per ton hot metal (50 lbs/net THM) and a production gain of 15% could be obtained as compared with a full sinter practice. With some pellets it became clear, that due to properties, commonly referred to as swelling, we got very poor furnace performances; so in studying the production of pellets, pellet quality should be regarded as the most important aspect of the problem.

Jan 1, 1972

By Ken Pill, Jennifer Nyman, Liz Moran, Keaton Belli, Chelsea Bokman

The Moab Uranium Mill Tailings Remedial Action (UMTRA) Project Site (Site) is a former uranium-ore processing facility near Moab, Utah, where production ended in 1984. The United States Department of Energy Office of Environmental Management (DOE-EM) is responsible for remediating the site to meet 40 Code of Federal Regulations (CFR) 192 standards promulgated by the United States Environmental Protection Agency under the authority of the Nuclear Regulatory Commission (NRC). Tailings from the former mill were stored in an unlined impoundment (tailings pile), totaling 16 million tons of waste material. Surface cleanup has been ongoing since 2009 and consists of removal and transport of tailings via rail to an off-site disposal facility in Crescent Junction, Utah, with an estimated completion date around 2029. Gravity-driven migration of tailings fluids and recharge through the tailings pile and former on-site ore storage areas in the northeastern part of the site led to contamination of site soil and groundwater. The primary contaminant of concern is ammonia, which was present in fluids used in the milling process (DOE 2003). Uranium is a contaminant of potential concern. An interim groundwater remedy is currently in place that utilizes a system of extraction and injection wells to limit discharge of contaminated groundwater into the nearby Colorado River. Exceedance of surface water standards for ammonia are a principal concern due to potential endangered species fish habitat areas along the Colorado River adjacent to the Site. When necessary, surface water diversion is used to lower surface water ammonia concentrations when suitable habitats develop for endangered young-of-year fish. Once surface cleanup concludes and the tailings pile is removed, a formal groundwater remedy must be selected and documented in a Groundwater Compliance Action Plan (GCAP). Under Title 1 of the Uranium Mill Tailings Radiation Control Act, natural flushing (i.e., monitored natural attenuation), may be selected as the groundwater remedy if it can be demonstrated that the remedy will achieve groundwater standards within 100 years. The 40 CFR 192 groundwater standard for uranium is 44 micrograms per liter (μg/L), and the target groundwater concentration of ammonia based on adverse effects on local fish species is 3 milligrams per liter (mg/L). A collaboration with Geosyntec Consultants, Inc. (Geosyntec), the Moab UMTRA Project team, the DOE Office of Legacy Management, the DOE-EM, and the Network of National Laboratories for Environmental Management and Stewardship was held in 2022 with the goal to generate actionable recommendations to develop a GCAP for the Site. The primary recommendation was development of a groundwater fate and transport model to predict the remedial timeframe under various groundwater remedies. This paper presents the results of soil and groundwater sampling, geochemical modelling, and laboratory column and batch reactor testing that will be used to parameterize a numerical model for estimating the remedial timeframe under various scenarios, including natural flushing.

Feb 1, 2025

The 2012 North American Tunneling (NAT) conference is expected to attract more than 800 tunneling and underground construction professionals. But don?t let the North American part of the meeting?s title fool you. Many of the attendees will come from all over the world. The meeting is scheduled for June 24-27 at the JW Marriott Indianapolis Hotel in Indianapolis, IN. In addition to two and a half days of technical programming, the accompanying exhibit has attracted 116 exhibitors in 140 booths. And the conference will include short courses, a field trip and plenty of social activities that will give attendees the opportunity to catch up with each other.

Jan 1, 2012

By H. W. Breathitt

World demand for phosphate rock has increased significantly in the past fifteen years. Prior to the early 1960's, most of the world production was obtained from Florida, North Africa, and Kola. As the 1960's arrived, new mines in North Florida and Eastern North Carolina, as well as expanded production in the Western United States, .increased domestic production. Many new mines overseas have also become significant. This includes countries such as Senegal, Togo, Israel, Spanish Sahara, and Jordan. Limited production, often as by-products, has been noted in South Africa, Brazil, and other countries. In recent years, and even recent months, we have learned of many new areas being considered as potential sources of phosphate rock. This includes Australia, Saudi Arabia, Mexico, Colombia, and Brazil. Many other developing countries have discoveries and are considering their new deposits, often with difficult beneficiation problems, as potential sources of phosphate. These factors all relate to expanded world, population and the demand for more fertilizer products to produce food. In this paper, four areas of the world with different types of phosphate deposits are briefly reviewed to contrast the ore beneficiation problems to Florida procedures--long considered the "standard" of the industry, at least for most of the United States' domestic production. Agrico has three large operating mines in Florida with total production capacity exceeding seven million tons per year. Here, pebble phosphate (+14 mesh material) is produced without the use of flotation, whereas all of the material finer than 14 mesh is subjected to flotation beneficiation to upgrade into high grade concentrates.

Jan 1, 1977

By D. Tahija

"Hecla's Greens Creek silver-gold-lead-zinc mill began using carbon dioxide rather than sulphuric acid to control pH in April, 2015. High pH levels in the grinding and lead flotation circuits are caused by the presence of re-mined cemented paste backfill in mill feed and need to be lowered to achieve good flotation response. Carbon dioxide has proven to be quite effective at controlling pH in the lead circuit with greatly improved pH target attainment. The allowable amount of carbon dioxide used is not limited by effects on lead-zinc selectivity and by poor froth characteristics, as was the case with sulphuric acid. Startup and implementation went smoothly, with few operational issues needing to be resolved. Early results from implementation of the system and its effects on mill performance will be presented. INTRODUCTION Hecla Greens Creek is a silver-gold-lead-zinc mine producing flotation lead and zinc concentrates, as well as a gravity gold concentrate. Besides their base metal contents, both the lead and zinc concentrates carry precious metals. The circuit is largely conventional, with SAG and ball mill grinding being followed by graphite removal preflotation, then conventional lead flotation and conventional zinc flotation. Both rougher concentrates are treated in regrind steps, followed by cleaner flotation. The cleaners tails are treated in an unorthodox bulk circuit to produce a low-grade zinc-lead concentrate marketed to Imperial Process smelters (1). BACKGROUND Backfill and Flotation pH The Greens Creek mine is an underground operation in which mined areas, after ore extraction, are backfilled, using rammed paste or pumped slurry. The backfill consists of mill tailings to which approximately five percent Portland or similar cement is added to enhance its compressive strength. Later mining near backfilled areas inevitably results in a considerable amount of old backfill being mixed with new ore before being shipped to the mill. Depending on orebody geometry, mining methods etc., the amount of backfill in mill feed is often 20% to 25% and sometimes even higher. Commercial cement, and consequently backfill prepared using it, contains excess lime. Because of this, backfill in mill feed is indistinguishable from the addition of a large amount of lime, so mill natural pHs often range above 10 or even 11. Similar situations have been recognized for many years at mines throughout the world (2)."

Jan 1, 2016

By N. J. Ferguson

An upsurge in demand for Australian coal has seen a much needed programme for a new generation of coal ports. This paper discusses design concepts in planning a large modern terminal, with the 25 Nt/a Port Kembla and the 50 Mt/a Newcastle facilities as practical examples. The need is emphasised for a studied Master Plan, in-built materials handling flexibility and staged development. Trends in bulk ship design and loading methods are noted. Features of these Australian designs include methods of dealing with poor ground conditions, functional civil engineering for environmental protection and optimal marine structure designs.

Jan 1, 1982

By Edwarde R. May

The Flambeau volcanogenic massive sulfide deposit located in northwestern Wisconsin occurs within the southern lobe of the Canadian Precambrian Shield. The steeply dipping tabular body occurs in a complex pile of cyclic schistose intermediate to felsic fragmental volcanics. No chloritic alteration pipe is recognized but adjacent rocks have undergone magnesia metasomatism and subsequently were metamorphosed to the upper greenschist and lower amphibolite facies. Although modest in size the deposit is of special interest in that it has an extensive and well preserved supergene enrichment blanket. Economic minerals found within a pyritized quartz-sericite schist are chalcocite and bornite overlying primary chalcopyrite.

Jan 1, 1976

By R. L. Miller

"Why should we care about diesel quality amidst the complexity of mining for minerals? Mining companies use a large amount of diesel fuel in their worldwide operations, both for moving ore from the mine to the processing mills or transportation hubs, and to generate power for onsite processing. This creates a large carbon footprint and poses both environmental risk with the production of emissions, and public perception of how those companies do their business as stewards of the environment while incurring a huge fuel expenditure. To improve diesel quality and create incremental value and greater environmental stewardship, the opportunity is broken down into several segments: 1. Mining industry diesel consumption 2. Fuel quality issues 3. Impacts on operations, economics, and emissions 4. Potential solutions 5. How improved fuel quality saves fuel expenditures 6. Economics of improved fuel quality 7. Recommended testing process 8. Fuel quality analysis checklist MINING INDUSTRY DIESEL CONSUMPTION The average grade of mined Australian ore bodies has halved, while the waste removed to access the minerals has more than doubled over the last 30 years, thus increasing energy consumption by 70%, (http://eex.gov.au/industry-sectors/mining/). Energy consumption and intensity in mining and mineral processing is rising at around 6% per annum. The mining industry is a huge consumer of diesel fuel worldwide, and the SRS Rocco Report of 4 August 2014 gives a better picture of recent trends. To better understand the reasons to reduce both fuel costs and consumption, we need to compare metrics on the amount of diesel consumed per each ounce of gold produced over the last decade. The rate of diesel consumption in gallons per each ounce of gold produced for the top five worldwide gold producers has doubled in the last 8 years from 12.7 gallons per ounce in 2005 to 25.8 in 2013 as shown in Figure 1 below.This is caused by 1) lower ore grades requiring more hauls to get same amount of finished gold product, and 2) haul trucks moving ore from deeper sources in open pit and underground mines. For one large multi-national gold mining company, the 2005 waste rock / production ratio was 52 metric tons/gold ounce; by 2013 that ratio had grown to 113 metric tons/gold ounce per the same SRSrocco 4 August 2014 report. Using the metrics above, the top 5 gold miners used over 591 million gallons of diesel fuel in 2013. The cost of diesel in dollars per ounce of gold has increased from $30.48 in 2005 to $101.14 in 2013, an annual appreciation of 16.2% as shown in Figure 2 below."

Jan 1, 2017

By B. Y. Wong

In April 1989, Southern Company Services signed a 3.5 year contract with DOE-PETC for generic engineering development of an advanced physical fine coal cleaning technology known as selective agglomeration. This technology offers the potential for significantly higher pyritic sulfur removal (90% +) than conventional coal cleaning (30 to 50%) while maintaining high Btu recovery (90% +). Engineering development will include bench-scale process development; component development; adaption or modification of existing unit operations; proof-of-concept (POC) module design, fabrication, testing and evaluation; and small commercial facility conceptual design. The information obtained during POC operation will be sufficient to result in a technical and economic design base that can support construction and operation of a small commercial plant. The major goal of the research is to develop selective agglomeration technology to achieve 90% pyritic sulfur removal while operating at Btu recoveries above 90% based on run-of-mine quality coal. This objective will be accomplished in two phases: Phase 1 -bench-scale and component development, and Phase 2 -POC module operation and conceptual design of a small commercial plant. Currently, work is proceeding under Phase 1 (Tasks 1-7) of the project. A review of the status of Tasks 1-7 is presented in this paper.

Jan 1, 1991

Flotation of dolomite from apatite was investigated, using dodecylamine hydrochloride or sodium oleate as the collector, in the absence and presence of inorganic modifiers such as sodium chloride. Separation of dolomite from apatite was anticipated from the single mineral experiments under various pH and collector concentration conditions. Flotation selectivity in the mixtures, however, was found to be limited. Attempts to control the flotation response of these minerals by addition of salts such as NaCI and KCI indicated that selectivity is possible. The presence of sodium chloride was determined to depress apatite flotation using sodium oleate leaving dolomite flotation unaffected (Moudgil and Ince, 1989). In this paper, conditions for optimum separation and the effect of major parameters on the desired selectivity are discussed.

Jan 1, 1990

By S. Bhattacharaya, S. Aich, B. K. Nandi

"At present scenario, there has been an increasing demand in the usage of blend coal in power sector; regardless of how, specific knowledge on burning and ignition characteristics and their interactive effects is stingy. In this work a detailed investigation was therefore taken up to study in detail the combustion characteristics of selected coal samples and their blends. Certain parameters were believed to have significant effects on combustion characteristics of individual coal and their blend. Major parameters identified for this purpose included proximate analysis, combustion conditions, heating rate, reaction regions, blending ratios, ignition condition, reactivity (or reaction rate) and burnout temperatures. With the help of these parameters, thermal features and reaction kinetics of the individual and blended samples were investigated through a qualitative and quantitative thermo-gravimetric analysis of two different individual coals and their blend carried out in an oxygen atmosphere from ambient to 800 0C at a constant heating rate of 10 K/min. A significant difference in terms of rate of mass loss with temperature correlated from Arrhenius expression along with coal-coal blend interactions and deviation from their expected behaviour, have been observed between the kinetics of individual and blend samples, the latter showing better ignition performance and comprehensive combustion phase. It was also found that particles having the lowest percentages of volatile matter and ash could possibly be blended with particles having high ash and volatile matter to meet the required burning condition which can be effective in terms of cost savings, improved fuel economy, reduced undesirable exhaust emissions and improved safety of coal burning equipment. Experimental results further indicated that individual samples and coal blend samples have different combustion features in terms of residue left, peak temperatures and activation energies. Thermodynamic activation parameters such as activation energy, pre-exponential factor and order of reaction were computed from the thermal data using the modified form of non-isothermal methods. INTRODUCTION Coal has been the most significant source of energy for India and would remain so at least for the next few decades. About 60-65% of the electricity produced in India is from coal fired power plants. Indian coal is of drift origin and of high ash content. Ash content of the coal burnt in power plants in India is in the range of 30-45%, most of it being Run-of-Mine coal. Most of the power plants in India source their coal from number of mines, most common being multiple sourcing. Usually coal from each source [1] is burnt separately. For various reasons mostly associated with logistics, burning of blended coal is rarely practiced. Blending of coal is practised mainly for improving the quality of available coal and meets the shortage of coal supply. Blending technique is generally practised in power industries to reduce the cost of power generation along with mitigating combustion performance or slagging problem ensuring consistency in as fired coal quality and also for environmental compliance. Blending good quality imported coal can improve coal quality and helps in preserving indigenous coal resources for longer period [2]. Use of blended coal shall reduce load on all boiler components (Mill, CHP, Boiler, AHP) & reduce O&M cost and increase the availability. The main purpose of blending is to get a blended as fired coal having ash, volatile matter and gross calorific value close to design coal and the two coals are to be blended in such ratio proportions that resultant coal blend have desired ash, volatile matter and gross calorific value [3]. For a coal to be blended certain parameters have to be considered to get compatibility with respect to combustibility ( Burn ability) of coal, grindability ( HGI), ash fusion behavior, swelling characteristics of coal, need to be checked"

Jan 1, 2017

By J. T. YIi-Penttiia

Dust formation and reaction mechanisms of copper matte particles were investigated in simulated flash oxidation conditions. The granulated and ground matte consisted of 72 wt% Cu, 4 wt% Fe and 21 wt% S. Experiments were conducted by using a laminar-flow furnace with a reaction gas composition ranging from 20 vol% to 75 vol% O2 in nitrogen. The particle feeding was carried out by the single-particle method and samples were quenched after a specified reaction time (approximately 0.05-004 s) in a cold water film. Experimental temperatures were between 800 and 1300°C. Chemical analysis was used to determine oxidation degree, and an optical and scanning electron microscope combined with energy dispersive analysis was used for characterization of the reacted particles.

Jan 1, 1998

By R. F. Korcak

There exists a wide array of industrial by-products that can be shown to produce beneficial responses in the soil/plant environment. These materials range from clean coal technology residues produced in the generation of electric power, particularly those high in gypsum, to mineral 'fines' produced by the aggregates and crushed stone industries. A number of these materials will be examined in relation to potential nutritional interactions in soils and crops with emphasis on trace elements. The significance of utilizing these materials in agriculture as a component of sustainable agricultural systems will be discussed.

Jan 1, 1995

By Thomas E. Finch

Engineering mine design encompasses a range of situations, and each situation has its own range of detail and fineness. The most time consuming, and perhaps repetitive design situation involves evaluation of new or alternate mining methods. Many man years are consumed examing alternatives conjured up at all levels of the design team, from the number crunching engineer to the engineering manager. Many alternatives are intuitively disregarded, some are briefly evaluated and thrown out, and many others are started but not completed due to time constraints. All alternatives have similar input characteristics and evaluation procedures. In the design process, production requirements are known and equipment production capabilities are generated to determine the numbers and costs of an equipment type required to match a given situation. The end result is some comparative figure, typically cost per ton or yard upon which a decision is built. The similarity of input data and problem solving techniques lends design to computer solution. Computer aided equipment evaluation is not new to the mining industry. An existing and accepted technique is exemplified with the programs used by truck manufacturers. Given a haulage profile and production requirements, the number of anyone capacity, horsepower, tire size, gear ratio or similar characteristically identified truck can be determined. This information can be outputted in varying degrees of complexity and the results used with equally varying percentages of credibility by design or production engineers. Most engineers are willing to accept-the computer's effort with relief and a desire for belief. Although inherently more complex, the ultimate end of computer manipulation should be total mine design. This is an extremely large step in the evolution of mine design, and not many engineers, including the authors, believe the results would be directly applicable or necessarily correct. As with computer simulations of an operation, total computer design of mines requires such complex and lengthy input that each mine might be manually designed more quickly and cheaply. The existence of computer programs which initially evaluate a mining system or part of the system is upon us, however.

Jan 1, 1978

By Charles F. Herbert

It is difficult to decide who most deserves credit for discovery of a mining district that has produced over 19.8 Mg (7,000,000 fine oz) of gold and still contains a large reserve of low grade ore. Fred Culver may have made the first discovery. In the summer of 1867, the ship's log of the Hudson's Bay Co.'s steamer Otter stated that Fred Culver had been found nearly lifeless from wounds, without provisions, but with some 2.3 kg (80 oz) of rough, placer gold in the bottom of his canoe; he told of finding a rich creek and an attack by Indians that caused the death of his two companions. The place of rescue was about 24 km (15 miles) southeast from the present site of Juneau. Possibly John Muir, the hero of modem environmentalists, should be given important credit. In 1879 he reported finding mineralization along what later became to be known as the Juneau Gold Belt, and it was his report that induced the cantankerous German mining engineer, George Pilz, to grub- stake some of the prospectors who had drifted into Sitka from the declining gold placers of Cassiar, British Columbia. Pilz also offered rewards to Indians who might report gold discoveries.

Jan 1, 1990