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By James W. Stewart, A. C. Colvine

Introduction Precambrian shield terrains the world over are experiencing an unprecedented level of concerted exploration for lode gold deposits. Reasons for this include: • The increase in the price of gold (following 40 years at $1.13/g or $35 per oz) has transformed targets that were once unattractive into viable exploration propositions. • The relatively suppressed market for other metals produced from shields (base metals, nickel, iron, and uranium) has further focused exploration attention on gold. • The preeminence of gold production from rocks of early Precambrian age, as shown in Fig. 1. • The relatively high grade of gold in these mines (averaging about 10 Wt or 0.29 oz per st) permits profitable extraction even under depressed gold prices. A number of these mines have been in continuous operation for more than 50 years. • The recent discovery of the major Hemlo deposit, in well exposed rocks adjacent to the Trans Canada Highway, has had a world-wide impact. This demonstrates that a major new deposit can be found even in an accessible, well exposed area of a shield; gold mineralization is susceptible to rational exploration; and new target types for gold exploration exist. This paper reviews exploration activity for gold in shield terrains worldwide (Fig. 2), excluding Early Proterozoic paleoplacers of the Witwatersrand type. Emphasis is on the Archean, which is traditionally the most productive part of the Precambrian shields. There is also an emphasis on the Canadian shield (Fig. 3) because of its vast areal extent and exceptionally broad range of mineralization styles. Geological Approach A basic problem exists in gold exploration when compared to exploration for base metal massive sulfides that predominated during the 1960s and 1970s. Gold concentrations cannot usually be detected by airborne geophysical instruments. The majority of massive sulfide deposits were detected using airborne electromagnetic detectors and magnetometers. A full 75% of Canadian gold discoveries can be credited to surface prospecting. The percentage is even higher in other countries. This observation allows for great optimism for two reasons: • Large areas of shield terrain are obscured by either surficial deposits (glacial tills, clays, and desert sands) or deep surface weathering. This makes them inaccessible to prospectors. Logic dictates that such obscured areas should have the same potential for gold concentration as has exposed bedrock. • The Hemlo style of mineralization, a roughly tabular ore body where gold is associated with disseminated sulfides, is different from the traditional prospecting target. The gold is hosted by discrete quartz veins. While it is true that limited discoveries were made at Hemlo more than 30 years ago, and this style of mineralization is present elsewhere, recent

Jan 12, 1984

By E. C. Robertson

D.G. Wilder I found the suggestion that the amount of displacement of a fault can be numerically related to the thickness of gouge or breccia to be both intuitively satisfying and intriguing. I have long agreed that there is some type of relationship between the amount of gouge and the amount of displacement of faults. I congratulate the author for developing a numerical relationship between them. However, I am concerned that the limits for applying this relationship be fully understood. An underlying assumption in this approach is that there is either a uniform thickness of gouge or breccia along a given fault or the thickness does not vary widely. Since it is not always possible to confirm this, the displacements derived by this method should be viewed with caution unless significant fault extent can be observed. At the Nevada Test Site, in drifts constructed in granite for test emplacement of spent nuclear reactor fuel, we found a fault with 0.3 to 0.4 m (12 to 16 in.) of clay gouge. Within a few meters of this location, the fault had no clay gouge, but rather consisted of a highly fractured zone with significantly altered rock and some slickensides. Based on Fig. 1, the 0.3 to 0.4 m (12 to 16 in.) thickness of gouge would indicate a displacement in excess of 30 m (98 ft). However, no gouge thickness would indicate essentially no displacement. Based on a quartz vein that terminated on the fault, and is not identified nearby, an estimated displacement of more than a few meters was made. This estimate is consistent with that obtained using the regression line proposed in the paper if the 0.3 to 0.4 m (12 to 16 in.) thickness for the gouge is used. However, using the regression curves with zero thickness would not yield results consistent with what was observed in the field. Therefore, it is important to recognize that the suggested procedure would properly yield a range of probable displacements. ? *Work performed under the auspices of the US Department of Energy by the Lawrence Livermore National Laboratory under Contract W-7405-Eng-48. Reply by E.C. Robertson It is certainly true that the t (thickness) of gg-bx (gouge and breccia) on a fault does vary along the fault. My observations have been that near the termination of a fault, the displacement d is small and the t is also small, whereas the maximum d and t will usually be found in the central part of the fault. The information on gg-bx and t of the fault found in granite in the NTS tunnel by Mr. Wilder could be interpreted somewhat differently than he does. He speaks of the fault changing within a few meters from 0.3 to 0.4 m (12 to 16 in.) of clay gg to "a highly fractured zone with significantly altered rock and some slickensides," but no gg. The highly fractured rock may be taken to be bx, rock not so finely ground as gg but still crushed by the fault movement, equivalent to the gg in my usage, and probably occupying about the same t. Mr. Wilder's estimates for the fault in the NTS tunnel for t of 0.3 to 0.4 m (12 to 16 in.) and for d of a quartz vein, in excess of "a few meters," would place the point on the low side of the central trend line in my Fig. 1, at the lower limit. There is, of course, a problem with determining d using displacement of only one planar surface. It would be greater or lesser depending on the rake of the movement. Finally, estimating the d of a fault from its t should be made with awareness of our present uncertainties, as pointed out by Mr. Wilder. Although the central trend line in my Fig. 1 has a ratio of d/t of 100, I have put the limiting ratios at 10 and 1000. Understanding of the values of the ratio will be improved only with collection of more data, for which the discussion of Mr. Wilder is much appreciated. G.C. Waterman E.C. Robertson's paper provides significant information to a geologist attempting to deduce fault offset by noting the products of structural dislocation. However, considerable mapping in underground and open-pit mines, and examination of structures produced in different geological settings, have convinced me that gouge and breccia thickness are controlled by geological conditions and fault movement. The following paragraphs suggest geological variables that control them. 1. Depth of Loading A near-surface fault resulting from tensional stress has more breccia/gouge than is produced by a similar stress at considerable depth. A deep-loaded compressional stress may produce a linear zone of schist, or structural dislocation may occur along an earlier formed belt of schist. Such "shear zones" are common in Canadian mines in precambrian rocks. In neither case can offset be directly deduced by an analysis of the minimal gouge/breccia in the shistose rocks. At greater depth, stress may be partially to wholly relieved by flowage. I vividly recall first noting the regional "Midas Thrust" in the Lark mine, Bingham Mining District, UT (where we called the structure the North Fault). My recorded notes, as I remember them, showed a narrow gouge streak separated the `Jordan" and "Commercial" limestone units from impure, muddy limestone beds of uncertain stratigraphic position. The visible structure did not indicate the great importance of this premineral fault

Jan 12, 1984

By Richard L. Nielsen

Early Models In 1906, Daniel Jackling demonstrated that relatively low-grade disseminated copper deposits could be mined profitably using mass mining technology. The ensuing years saw a gradual increase in activity aimed at defining ore targets for large reserves of disseminated supergene enriched copper ores. Prominent organizations, such as Calumet and Hecla, Phelps Dodge, and Asarco, established teams of geologists. They focused on interpreting oxidized and leached outcrops over supergene chalcocite ore bodies. Techniques to evaluate oxidized cappings and predict the presence of underlying supergene chalcocite enrichment ores were developed and put to practical use by the late 1920s and early 1930s (Locke, 1926; Blanchard 1939, 1968). Oxidized capping evaluation centered on recognizing oxidation products derived from the various sulfide minerals found in disseminated deposits. Chalcopyrite, for example, oxidizes to a characteristic reddish pitch limonite that exhibits characteristic boxworks. Chalcocite is recognized by earthy, powdery, indigenous, hematite-rich limonite. Pyrite weathers to characteristic cavities and yellowish jarositic limonites. Evaluation of cappings is a systematic effort of noting and estimating the preoxidation sulfide mineral content and copper grade in an oxidized outcrop using mineral composition and morphology of oxidation products. Several complications were noted early in this technology's development. For example, the associated pyrite content of protore greatly influences mobility of copper and iron during oxidation. High-pyrite disseminated sulfide assemblages, upon oxidation, mobilize iron as well as copper. This results in oxidized rock with many leached cavities, poorly developed boxworks, and poorly developed indigenous limonites. This effect was noted and used in a general way to identify possible development of supergene chalcocite. Indeed, strongly leached outcrops could be present over enriched ore. A second complication is the manner in which associated gangue minerals influence the products of sulfide oxidation. Reactive gangue minerals, such as carbonates, neutralize acids formed by sulfide oxidation. Thus, copper cannot be moved from oxidized outcrops. Relatively nonreactive gangue, such as quartz-sericite-altered schist or porphyry, neutralizes supergene acid very slowly. This allows copper to move out of the oxidized outcrop. Recognizing these features was closely allied to an early understanding of the processes involved in oxidation supergene enrichment of disseminated ores. Field observations and mineralogic mapping of oxidized cappings were important in discovering many important porphyry copper ore bodies during the 1920s. These include Silver Bell, Miami-Inspiration, and Morenci in Arizona; Tyrone, NM; and Ely, NV. Capping interpretation like-wise led to discoveries at Mineral Park and Esperanza in Arizona during the exploration surge that accompanied high copper prices in the 1950s. In the 1960s, Kennecott research geologists refined and quantified some aspects of oxide capping appraisal (Anderson, 1982). The ratio of jarosite plus hematite to the total limonite assemblage was shown to be proportional to amount of copper leached from the oxidized outcrop. Thus, if the limonite assemblege in an oxidized capping consists of 40% goethite, 20% jarosite, and 40% hematite, 60% of the original copper is leached. The amount of copper remaining in the capping, together with the limonite mineralogy, can then be used to estimate original copper grade in the outcrop, amount of leaching, and chalcocite enrichment grade (in feet/% copper) in the underlying supergene chalcocite zone (Anderson, 1982). Evaluation of oxidized cappings and prediction of chalcocite blanket ore targets were successfully used in the western US and South America through the 1950s. Discovery of new porphyry copper provinces in other regions during the 1950s and 1960s, however, underscored some limitations of capping evaluation technology. Porphyry copper systems in Australia generally oxidize and weather to hematitic cappings. Presumably, the semiarid monsoon environment and the long time that the sulfide systems have been exposed to oxidation on that continent lead to development of relatively stable hematite. It is now recognized that jarosite and goethite persist metastably in the limonite assemblage in all areas once sulfide oxidation is complete and acid generation ceases (Brown, 1971). The assemblage eventually converts to hematite. But, climatic conditions in arid southwest US allows metastable limonites to persist for a long time, thus providing information on former sulfide assemblages. Another problem occurs when applying capping evaluation techniques in extremely humid re-

Jan 12, 1984

By G. Kemp Williams, Peter H. Dohms, Barry A. Hillman, Laurie A. Williams

Introduction In 1969, the US devalued the dollar. This action, over the ensuing four years, resulted in the deregulation of the price of gold. It also resulted in the rebirth of the US gold mining industry. The extreme inflation of the 1970s led to a dramatic increase in the price of precious metals. Consequently an increased interest in the exploration for and mining of gold has occurred. The enthusiasm began among the few existing producers who had survived on marginal profits during a prolonged period of fixed prices and constantly rising costs. Then came the newcomers. They saw the opportunity to enter an industry that historically had been characterized by small operations with low capital costs. It was not until the late 1970s that large mining companies recognized that investment in gold projects could generate cash flows that previously were associated only with world class deposits. Today, gold mining has tremendous appeal for the entire mining community, despite the dramatic price drop from 1980-81 levels. There are even major firms with no mining experience attempting to gain a foothold. But how easy is it to enter the US gold mining industry? What are the basic factors that influence the economics, and what are they likely to be for the remainder of the decade? How has the industry changed, and what is the composition of the active companies? An examination of the industry will help to answer these questions. US Gold Mining History The current gold rush is actually the fifth in the 208-year history of the US. The first recorded commercial mining of gold occurred in the Southeast in 1799. It produced from small, yet relatively rich placer deposits. The famous California gold rush of 1848 placed the US as a major producer with 62.2 t (2 million oz) of output by 1850. More than 93.3 t (3 million oz) was mined by 1853, equivalent to 63% of world production. But the richest placers were soon depleted. By the end of the Civil War, annual production had declined to well below 62.2 t (2 million oz). The government then relaxed prices and the industry again began to recover. The 1870-80s saw the great mines of the California Mother Lode and Grass Valley brought into production. This was followed by development of Nevada's Comstock, which later became one of the world's largest producers. Discovery of gold at Cripple Creek in 1892, coupled with technological advances in the use of large dredges in the California placers, resulted in significant production increases. And, with the expansion of the Homestake mine at Lead, SD, in 1898, US output again surpassed the 93.3-t (3 million-oz) level. By 1905, produc¬tion climbed to 124.4 t (4 million oz) with the addition of the Alaskan deposits, new output at Goldfields, NV, and the introduction of the cyanidation process. During World War I, the government forced the shutdown of gold mines. By 1920, total production had dropped to little more than 62.2 t (2 million oz). Many mines remained closed until the Great Depression. This, combined with the devaluation of the US dollar in 1934, resulted in a gold price increase to about $1.13/g ($35 per oz). Then began the fourth gold rush. By 1937, output had again surpassed 124.4 t (4 million oz). This was due to an increase in primary production and new byproduct production from base metal operations. Unlike the previous period of growth, however, this period was short-lived. The imposition of Executive Order L-208 in 1942 once again forced the shutdown of most mines. At the end of World War II, cumulative US gold production had

Jan 12, 1984

By James W. Babcock

Introduction Geologic ore deposit models are key to any exploration program. In the search for ore deposits, the geologist first tries to understand the genesis of known ore bodies. During reconnaissance and evaluation of a prospect, a valid exploration model of a known ore body can help the explorationist focus on critical geologic attributes of the target area. Exploration models can conserve exploration funds that might otherwise be spent collecting data not critical to delineating an ore body. This article examines the benefits of using models, factors to be considered in developing models, and examples of the way recent models have evolved. New exploration concepts and techniques are developed in response to increased competition in the mining industry. New techniques can give an exploration group a slight edge that may translate into a discovery. Mine geology staffs may find previously unlooked for extensions to the ore body by applying new genetic concepts. When all appropriate techniques are applied to exploration and mining efforts, greater efficiency and increased profits may result. Before the 1950s, scientists speculated on the genesis of ore deposits and other natural phenomena. But they presented their results as if universal and final, even though they applied to only a few deposits. Because it did not provide sufficient criteria to distinguish the genetic processes, the classic conflict between hydrothermalists, epitomized by Lindgren, and nonhydrothermalists did not help the minerals explorationist much. Important information about individual ore deposits, such as age of the host rocks and mineralization, composition of the mineralizing fluids, and separation of mineralizing and nonmineralizing events in the paragenetic sequence, could not be gathered with past equipment and methods. But these early concepts of ore genesis contained excellent empirical observations, perhaps the most useful parts of these hypotheses. Through the years, prospectors have found most of the ore deposits in the world using direct observations and their collective empirical wisdom - an empirical model. The geologist's contribution becomes increasingly important after the more obvious surficial ore deposits have been discovered. In Europe and North America that time has arrived. With the introduction of sophisticated equipment and matching techniques in the 1950s in geochronology, geochemistry, geophysics, remote sensing, and computers, many previously equivocal genetic questions can be resolved. This proliferation of highly technical data has allowed the construction of many well-founded genetic models.

Jan 12, 1984

By Fred D. Fox

Introduction Abandoned mill tailings deposits can be found throughout the northern Black Hills of South Dakota, indicative of mining activities that originated during the great gold rush days of 1875. The Refuse Act of 1899 attempted to regulate the deposit of debris from mines and stamp works. But, in most cases, huge volumes of mine and milling wastes were disposed of in ways no longer considered environmentally acceptable. Restoration of gold mill tailings deposits is a relatively new practice to the Black Hills area. This discussion will present an overview of the environmental setting, environmental problems associated with past tailings disposal practices, practices being implemented to solve these problems, and prospects related to present the future tailings disposal practices. Environmental Setting An understanding of climatic characteristics, geology, surface and ground water hydrology conditions, and the physical and chemical characteristics of selected mill tailings deposits will aid in understanding environmental problems associated with tailings disposal and restoration. Climate The Black Hills region is in the middle latitudes, which have a

Jan 11, 1984

By W. R. Bruce, R. W. Wittkopp, R. L. Parratt

Introduction The Relief Canyon gold deposit is about 24 km (15 miles) east of Lovelock at the south end of the Humboldt Range in northwestern Nevada. The deposit, is in the Relief-Antelope Springs mining district, which has historically produced silver, antimony, and mercury. There is, however, no mention in the literature of commercial gold production. Fluorite prospects at the gold deposit site have had no reported production. At Relief Canyon, the Late Triassic Grass Valley formation overlies and is in fault contact with the Late Triassic Natchez Pass formation. Epithermal disseminated gold mineralization is found within the various types of fault breccia between these two formations. Geology The Natchez Pass formation of Late Middle to Late Triassic age is composed of more than 300 m (985 ft) of massive gray to dark gray locally carbonaceous dolomitic limestone. Some minor beds of shale and siltstone up to 1 m (3 ft) thick are found in the project area. The limestone is locally silty or sandy. The color of this formation below the oxidation base ranges from gray to black and appears to be a function of carbon content. The Grass Valley formation of Late Triassic age is composed of more than 200 m (655 ft) of interbedded units of thinly parted argillite, hard gray to brown quartzite, siltstone, and shale. Within the oxidation zone, these units are olive gray. A few beds within this formation are slightly calcareous and a number of sections, especially those containing shale, are dolomitic. Below the oxidation zone, the quartzite beds are often slightly carbonaceous and the argillite, siltstone, and shale beds are often highly carbonaceous, giving them a black color. Two types of intrusive rocks have been recognized at the Relief Canyon deposit. Both appear to predate mineralization. Fine to moderately fine grained quartz monzonite dikes, up to 3 m (10 ft) thick, were encountered in several drill holes. In a number of intervals, these dikes have undergone either propylitic or argillic alteration. The age of these types of dikes is not known. It appears, however, that they are either Jurassic or Cretaceous. No gold mineralization has been found in this type of dike. Diabase dikes were also encountered in a number of drill holes. These dikes have almost always been propylitically altered. Although the exact age of the diabase dikes is not known, they are probably equivalent in age to the quartz monzonite dikes. Quaternary alluvium is found forming fans at the base of steep slopes and as recent fill in present day drainages. The alluvium is composed of either Natchez Pass limestone or Grass Valley quartzite and siltstone, depending on which unit served as the bedrock source. A significant portion of the Relief Canyon deposit is covered by Quaternary alluvium. Figure 1 shows a generalized geologic map of the Relief Canyon area. At the deposit's site, the Grass Valley formation appears to have been thrust over the Natchez Pass formation. The age of the thrust is probably correlatable with the Nevadan Orogeny, which gives it a Jurassic-Cretaceous age. The general strike of the thrust, referred to as the Relief Fault, is in a northwest direction. The strike of the bedding of both the Natchez Pass and Grass Valley formations roughly parallel the strike of the Relief Fault. The general dip of both the Natchez Pass and Grass Valley formations is in a southwest direction. The general dip of the Relief Fault, in the area of the Relief Canyon gold deposit, varies and has the appearance of a northeast-southeast striking anticline that plunges in a southwest direction. A small fold perpendicular to the plunge of this anticline forms a dome over the southerly portion of the Relief Canyon deposit. A number of northeast and northwest trending normal faults slightly offset the Relief Fault. Because of their small displacement, they are not shown on the generalized map. Gold Mineralization Gold mineralization occurs along the highly brecciated fault contact between the Natchez Pass and Grass Valley formations. Weak gold mineralization often occurs up to 2 m (6.5 ft) above the thrust in the Grass Valley formation. Most of the ore grade mineralization, however, is present below the Grass

Jan 11, 1984

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 Richard J. Stoehr

In his time on earth, man has mined about 93.3 kt (3 billion oz) of gold. About one-half of this has been produced in the past 35 years. Roughly one-third of this gold is held as reserve in Central Banks. One-third held is in personal reserves. And one-third is held as jewelry or other fabricated products. The world per capita availability of mined gold is increasing. There are many forces shaping gold's future. Gold currently has a favorable exchange ratio. This is encouraging many new gold mining developments. Production is increasing and new recovery techniques are surfacing. The use of gold in the financial markets is increasing in some unique ways. And gold is again fulfilling its role as a universal money, making its future predictable. Gold Today Man is currently mining gold at the rate of about 1.4 kt/a (44 million oz per year). This adds about 1.5% of newly mined gold to the total supply each year. This production plays a small role in the pricing mechanisms since much of the 93.3 kt (3 billion oz) mined is available to the market in some form. The 31.1 kt (1 billion oz) held by central and international banking

Jan 11, 1984

By Edward H. Jr. Boyle, Paul R. Thomas

Introduction It is estimated that nearly 40% of all the gold ever mined in free world countries has been produced by South African operations. There is concern that this immense productive capacity will soon decline. Such a decline could reduce world gold supply and lessen the ability of market economy countries (MEC) to exert the major influence upon the world gold market. This paper addresses three general issues concerning South African gold production: • How much gold remains to be recovered from the demonstrated resources of the 44 evaluated South African mines; • How long can production at current levels continue; and • What will be its cost. Table 1 provides production data on the world's 10 largest producers for the years 1971, 1981, and 1983. As shown, South African gold production declined significantly from 1971 through 1981. As of 1981, annual South African gold production had dropped by more than 311 t (10 million oz) in 10 years. Despite this, its mines still produced 656 t (21 million oz) that year, about half the world total. From 1981 through 1983, this trend was reversed. Annual South African gold production increased by more than 21.8 t (700,000 oz). The trend reversal was also evident in the US, Canada, and Australia. This illustrates the general, inverse relationship between gold price and gold production. In South Africa especially, as gold prices increased substantially during the 1970s, the average grade of ore mined (pay limit) was lowered as new, lower grade areas and materials were exploited. Gold production declined in volume. But gold revenues increased, and the lives of the operations were extended. The 1981-83 period was one of declining (US dollar-based, average annual) gold prices that resulted in increasing South African production as average ore grades were raised. South Africa currently accounts for 50% of world production (Fig. 1). Figure 2 illustrates South Africa's contribution to total MEC production in 1971 and 1983. Although South Africa's relative share of production has declined, it still represents the single most important source of new gold supply. In 1983, it supplied more than 65% of total MEC production. South Africa currently produces more than twice as much gold as the Soviet Union (ranked second in world production), and 11 times more gold than the US (ranked fourth in world production). Barring major discoveries, South African gold production is expected to peak in 1987 and begin to decline again throughout the 1990s. The rate of decline is ex-

Jan 11, 1984

By Clement K. Chase, Manuel R. Lastra

Introduction The choice of using a dump or heap leaching process or a fine grinding-agitation leaching process is based on the grade and tonnage of the ore body. Low grade ore and small tonnages are best treated by a low-cost method, such as dump or heap leaching. Permeability, however, is the critical factor in a successful leaching operation, regardless of tonnage or grade. Without permeability in the leach bed, no percolation, no dissolution, no extraction, and no success are possible. Solution delivery to the ore is also essential. Solute-to-mineral contact is a requisite. The solute must be effectively strong and not be depleted in active dissolvers, such as acid, ferric iron, cyanide, and oxygen. It must also be moving across the mineral to be dissolved. No pockets of stagnant solution can be allowed. A third factor is recovery of the now-dissolved metal in the pregnant liquor. All such liquor should be gathered at a reasonable metal concentration to make further pumping and processing worthwhile. Control and recovery of pregnant solutions are probably the most ignored factors of solution mining. This stage in recovery is historically where many operations either fail or become viable. Permeability A permeable heap is normally achieved by taking into account heap building/construction and agglomeration/pelletizing. Even though heap building is critical for good permeability, it will not be discussed here. It has been described in detail by others. Instead, this analysis will concentrate on agglomeration and pelletizing. Differences between agglomeration and pelletizing are emphasized. Agglomeration is referred to when small particles are being attached to larger particles -the ore presents a complete size distribution. On the other hand, pelletizing is a process where all particles are fine (less than 850)µm or 20 mesh), as with mill tailings. The difference is important from two points of view: • Mechanical devices used to achieve the desired attaching action; and • Binder dosages. Mechanical Devices Due to the absence of large particles that fines can become attached to, pelletizing should be done in a centrifugal-action-based device, such as an inclined rotating plate, a rotary drum, or a series of belt-transfer points. In agglomeration, large particles play a part as a natural nucleus for the attaching action of fines. Hence, a tumbling action as strong as for pelletizing is not required. Common methods to accomplish agglomeration include: • Cascade belt system with either all belts moving forward or some of them moving forward and at least one moving backward; • Combination of belt plus inclined vibrating chute; • Large concrete mixer drum; and • Belt discharging directly onto stockpiled ore. The proper agglomeration system will depend basically on the size distribution of the ore and the amount and type of clays present. In all of the agglomerating systems described, the tumbling action takes place at the transfer points either from belt to belt, or from belt to inclined chute ore belt to ore pile. Usually, two or three mixing bars at each transfer point are also included to help the mixing action. Plows to move the ore around on a moving belt are also used.

Jan 11, 1984

By Stanley W. Ivosevic

Introduction Increasing investor and industrial demand for gold is not being matched by new mine output from traditional sources. This forces the exploitation of alternative natural and industrial resources to supplement traditional sources. A traditional natural resource is either high- to medium-grade ore, of gold or silver with coproduct gold. Or, they are medium- to low-grade base metal ores having byproduct gold. Traditional ores are also frequently extracted at relatively great expense by selective underground mining. Alternative natural resources include low-grade, near surface ores and the products of mining-dumps - and of onsite processing - tailings. Traditional industrial resources are newly mined gold or that obtained from plant sweepings. Their alternative analogues include gold in old scrap and chemical sludges and precipitates, notably those incidental to the electronic industry. These alternative gold resources have several attractions. They are untapped and abundantly available. Having been overlooked by prior metals suppliers, title to many alternative resources is easily acquired. Or, the value of such a resource may go unnoticed until its gold is rendered exploitable by an advance in extractive technology or an other approach. This article addresses the effect of large tonnage, low-grade lode ores on gold supply. Exploiting these ores is rendered commercial by their amenability to bulk mining by modern large-scale mining and metallurgical operations requiring little selectivity. Placer gravels, perhaps the earliest type of gold ore mined, also are bulk minable. But these fall outside the definition of being lode deposits. Most lode bulk mining is from surface open-pit mines. Some, however, is from underground by such large-scale mining techniques as room and pillar, vertical crater retreat, and end slicing. The low-grade gold ore being discussed averages 2.8 g/t (0.082 oz per st). Output How greatly do bulk minable, low-grade resources effect supply? 1982 was a somewhat healthy year for gold mining and exploration in North America in spite of the general depression in the mining industry (Table 1). The wildly fluctuating price of gold bullion had stabilized at an annual aver¬age of $12.08/g ($376 per oz) - (Handy and Harmon base price). Half of the 1.3 kt (43 million oz) of 1982 world gold mined were from South Africa's high-cost, medium-grade, selective underground mines. Of the remaining half, 20%, or about 130 t (4.2 million oz), was mined in North America. This includes Canada and the US - the third and fifth largest gold producers in the world. Of that North American production, more than 62.2 t (2 million oz) of gold, or about 50%, came from bulk mining of gold ores with or without co-product silver. Most of this was in the US, where bulk minable gold-silver lode ore produced nearly 31 t (1 million oz). This amounted to 60% of the nearly 46.6 t (1.5 million oz) produced in the US during 1982. An additional nearly 7.7 t (250,000 oz) of gold were produced as the byproduct of bulk mining of cop¬per ore, for a total of 34 t (1.1 million oz) of gold, or 75% of US gold production by bulk mining. To further illustrate this, US placer, mine dump, and related operations produced an insignificant 4% of US gold output at the time. Exploitation of bulk minable gold deposits is becoming increasingly important worldwide. Most new Australian gold mining announcements are of bulk minable developments. This trend will increase in North American mining as more large tonnage, low-grade operations come onstream in Canada, where much current production is from underground. It will increase with the general climb in Mexican gold mining. And it will grow with expansion of the 12 t/a (400,000 oz per year), gold production of the Dominican Republic. Production Metal price and operating costs make these large tonnage, low-

Jan 11, 1984

By Jan Krason

Introduction The South African Witwatersrand is the world's largest gold-bearing province and gold producing district. But the production capacity of the very big mines (Vaal Reef, Driefontein Cons., Western Holdings, Western Deep Levels, and others) is not as large as that of the Soviet Union's Muruntau gold mine. The Muruntau gold-bearing ore body was discovered in 1956-57, and has been in operation since that time. In the early 1970s, this mine produced 70 to 80 t/a (2.3 to 2.6 million oz per year) of gold. Within the last few years it has become the world's largest producing gold mine. Russian authors do not publish ore reserve or rate of production data. But it is very likely that the Muruntau mining complex, which involves several ore bodies, is currently producing about 143 t/a (4.6 million oz per year) of gold. The author has confirmed this by personal information from several knowledgeable Russian geologists. This amounts to one-half of

Jan 11, 1984

By Ramon S. Pizarro, W. Joseph Schlitt

Introduction Trying to synthesize innovations occurring in gold processing is not simple. At present, this technology is exploding, brought about by the high price of gold. This price has compelled the mining industry, as a matter of survival, to produce gold. The frantic search for gold deposits continues, from glacial areas to tropical jungles. The search has been fruitful. New exploration techniques have helped to locate deposits. Some are amenable to conventional processing techniques. Others are in unusual settings that challenge the ingenuity of extractive metallurgists. Ore deposits containing traces of invisible gold, refractory mineralization, and impurities and deleterious materials that previously defied treatment are now being profitably processed. Metallurgists have developed various processing methods and equipment for gold ore treatment that reflect differences in ore characteristics, grade, mineralogy, reserves, topography, climate, and location. These processes can be categorized as: • For nonrefractory or readily leachable gold ores - vat leaching, heap leaching, and agitation leaching. • For ores containing free gold - gravity concentration methods, followed by mercury amalgamation, and • For refractory gold ores - flotation, roasting, pressure oxidation, followed by cyanidation leaching. In actual practice, a combination of these processes may be required. This is because gold mineralization, although usually simple, may be associated with other mineral species or with clays that make the ore difficult to treat. Furthermore, after the gold has been solubilized to form a cyanide complex, one of two important processes must be applied to recover the gold. These ore zinc cementation (Merrill-Crow process) and Carbon adsorption, desorption, and electrowinning. New techniques for leaching and recovery are also advancing. These will become the bases for the next generation of gold processing plants. Nonrefractory Gold Ores Nonrefractory ores with readily leachable gold are among the .easiest to treat. Even so, non-refractory ores from epithermal deposits are likely to contain too much clay. When these ores are damp, they may clog the screens and cone crushers. Depending on the type of process, scrubbing the crushed ore is very effective in overcoming clay problems. After dewatering, the slimes enter the cyanidation process and the thickener overflow is stripped of gold values and recycled. Semiautogenous (SAG) mills are also gaining popularity due to lower capital costs and ability to treat wet, sticky ores. If properly sized, the secondary crushing and screening plants, together with a secondary ball mill, can be eliminated. When the ore is of such a low grade that it can not support fine grinding, vat or heap 'teaching may be applicable. These systems are being actively applied in the US and elsewhere. Rapid construction and low capital costs are important features of these plants. Vat leaching is ideally suited for fast leaching ores and small tonnage deposits located in areas with harsh winters or wet tropical climates. For effective vat leaching, the ore is crushed to less than 25 mm (1 in.), agglomerated with lime or cement to control fines, and leached with sodium cyanide solution (upflow or downflow). The degree of size reduction is a tradeoff between increased costs of comminution and improved gold recovery from smaller particles. However, the size of material

Jan 11, 1984

By James L. Evers

Introduction There is a continuing search by government and industry for new mining methods that increase production, improve the working environment, reduce labor intensity, enhance recovery effectiveness, improve the reclamation process, streamline the flow of material, and reduce the unit cost. Hydraulic extraction and transport methods have outstanding potential in several of these areas. For example, hydraulic fragmentation produces a significant reduction in respirable dust. The inherent simplicity of hydraulic mining equipment and the resulting high reliability are recognized. Tool wear can be virtually eliminated. There is great flexibility in the size, shape, and orientation of the mineral deposit that can be mined. The cutting, gathering, and transport operation can be controlled remotely, providing more safety for workers. The threat of tool sparking and the resuiting dust or gas ignition are also eliminated. Although hydraulic mining was introduced in the US before the turn of the century, it has not gained the degree of acceptance found in other countries such as the Soviet Union, China, Germany, and Canada. The Hydraulic Mining Research program, initiated by the US Bureau of Mines in 1958, and the subsequent research by others has provided much of the basic technology required to spur the development of these methods into a viable extraction alternative. This article updates the development of hydraulic mining in the US. Particular attention is paid to monitor mining, the long-wall hydrominer, borehole mining, and the semi-autonomous mining system. Whenever the particular mining method is also being developed in other parts of the world, a brief description of the world-wide state of the art is given. Finally, allied studies designed to optimize these mining techniques are described. Hydraulic Monitor Mining Since 1958, the US Bureau of Mines (USBM), the Department of Energy (DOE), and other government agencies have funded more than a dozen programs to investigate the potential of hydraulic monitor mining. For this method, water is directed on the coal face from a monitor nozzle. Sufficient velocity is used to cause local fragmentation (Fig. 1). Coal particles are washed from the face to a trough. They are then flumed to a collection point. Oversized particles are directed through a breaker and then into the troughs. Monitor field tests were conducted in the 1960s in bituminous and anthracite coal at the following locations: • Rosylin No. 9 mine, WA (Nasiatka and Badda, 1963); • Rosylin No. 10 mine, WA (Price and Badda, 1965); • Sugar Notch mine, PA (Buch, 1964); • Thomson Creek mine, CO (Palowitch and Malinka, 1964); and • West Lebanon, PA (Fowkes and Wallace, 1968). In 1975, Ingersoll-Rand Research Inc. was awarded a two-year USBM contract to develop a continuous coal mining system using water jet fragmentation and

Jan 10, 1984

By William H. Dresher

Introduction A subject long debated but rarely resolved involves the interaction between the academic community and industry in the sponsorship and conduct of research. The subject is, unfortunately, too often treated as a controversy involving ethics and the profit motive. But, in fact, the two branches of society - industry and academia - are necessarily part of the same system. In the last 25 years, the realization has developed that research alone is not sufficient to bring about use of those results to benefit society through productivity gains and economic growth. The flow of events shown in Fig. 1 was generally believed to be the steps required for reaping the benefits of research. Today, research still has the very important role of creating new information or new ideas. However, it is recognized to be but one of many steps in a process that we now refer to as technological innovation. Charpie (1967) said, "Technological innovation is the driving force behind economic growth in the US. In this context, innovation means the process by which a new idea is successfully translated into economic impact within our own society, by providing better products, and simultaneously creating new jobs in the manufacturing of those products. Only after an investigation is put into sufficient use to have an economic effect, is it termed an innovation." One of the best depictions of the technological innovation process is shown in Figs. 2 and 3. These are adapted from Haeffner (1975), and modified by the Committee on Mineral Technology of the National Research Council (1981). Figure 2 identifies the many steps involved in the process. It designates a generalized form for interaction between the academic community and industry. Figure 3 identifies the steps in the "innovation activity" box of Fig. 2. Academia is generally responsible for developing of new knowledge and ideas, as well as maintaining and enlarging the collected body of scientific knowledge. Industry is responsible for developing new products, methods, and processes. Both institutions operate for the benefit of society, but under different driving forces. In academia, the driving force is the achievement of academic excellence. In industry, the driving

Jan 10, 1984

By T. W. Martin, E. M. Frizzell

Jan 10, 1984

By Vas P. Kenyen

Introduction Autogenous, semi-autogenous, and conventional grinding circuits are used to process ores. Current emphasis, though, is towards the use of the first two systems. The reason is the generally lower initial investment required for an autogenous or semi-autogenous circuit as compared to a conventional comminution circuit. However, the effect of operating costs that will significantly affect overall comminution economics is not considered in this evaluation. About 70% of the energy in a mineral processing plant is used in the comminution circuit. Until the 1974 oil embargo, energy was relatively inexpensive. Since that time, the cost of energy has escalated 40-fold. As a result, energy cost now forms a significant portion of operating costs in comminution. According to a survey conducted by General Electric Co. (Powers, 1982), the price of electricity will continue to rise 15% to 20% per year, in spite of the current oil glut. Both autogenous and semi-autogenous grinding circuits consume more energy than the conventional comminution circuit. Higher consumption ranging from 16% to 35% has been reported in the literature (Bassarear, 1982; Gjerdrum. 1979; Betts, 1982; and Christie, 1979). Consequently, even a small increase in energy costs is likely to have a significant impact on operating costs. Therefore, the economic decision of selecting an optimal comminution circuit is important. More than ever before, it is now critical to include in the evaluation of communition circuits not only the capital costs, but also the operating costs; in particular, the expected energy cost over the life of the operation. Failure to do this may lead to a less than optimal comminution circuit. Design A comminution system should be designed only after bench-scale or pilot plant testing of a representative sample. Conventional crushing and grinding circuits can be designed on the basis of the Bond Work Index (Rowland and Kjos, 1980). Over the years, sizing of full-scale rod and ball mills from Bond Work Index data has become well established. Autogenous and semi-autogenous mills cannot be sized based on bench-scale tests. Pilot plant tests are necessary to determine the amenability of the ore, critical size buildup, and power requirements. For this study, conventional and semi-autogenous grinding circuits were designed for ore throughput of 12, 24, 36, 48, 72, and 90 kt/d (13,200, 26,400, 39,600, 52,000, 79,200, and 99,000 stpd). Rod and ball mills were scaled up on the basis of the Bond Work Index. A two-stage semi-autogenous grinding circuit was designed on the basis of pilot plant test results. In designing a grinding system, emphasis was placed on operating flexibility. The largest commercially available equipment was

Jan 10, 1984

Introduction The frigid regions of Canada's high arctic islands are well-known for major oil and gas explorations. But mining leads the way in development. Cominco Ltd.'s Polaris mine is in its second full year of commercial production, and outstripping the rated production capacity foreseen by its planners. The Polaris mine is located on the shore of Little Cornwallis Island, near the settlement of Resolute Bay. The mine property is compact. It occupies about 3 km2 (0.5 sq miles) of the vast Arctic landscape. A number of mining and technological firsts were necessary to bring a mine into production at this 76° North latitude. Development constraints that had to be overcome included climate, remote location, and the limited ice-free shipping season. Construction of the entire process plant and ancillary facilities took place on a barge in the South where labor and supplies are plentiful. The plant was towed through ice-strewn waters to its permanent location at the mine site 1450 km (900 miles) from the North Pole. Freighters moved the first shipments to the South during ice break-up in late summer 1982. Most of the zinc concentrate is sold to European smelters. The remainder is tolled at a custom smelter in Europe, and the resulting metal is sold by Cominco. Lead concentrate is sold to European smelters. The mine was designed to produce 177 kt/a (195,000 stpy) of zinc concentrate and 41 kt/a (45,000 stpy) of lead concentrate. Actual production in 1983 was considerably greater, with both mill throughput and feed grades being higher than forecast. Production was 217 kt (239,300 st) of zinc concentrate and 51 kt (56,300 st) of lead concentrate. Average milling rate was 2.3 kt/d (2500 stpd), well above the designed capacity of 2.1 kt/d of (2300 stpd). Arctic Climate Extremes The climate is typical of the Arctic: very dry with short, cool summers and long, cold winters. From November through February, there is continuous darkness. Continuous daylight prevails from April to August. Precipitation is light, rivaling the Sahara desert at about 127 mm (5 in.) of rain and snow a year. An almost continual wind forms deep snow drifts against any obstacle on the ground. Temperatures range from a winter low of about -50°C (-58°F) to a relatively pleasant 15°C (59°F) during sunny days in August. Freeze-up begins in September and ice thickness increases from less than 0.3 m (1 ft) in early October to a maximum of about 2.1 m (7 ft) in May. The permanently frozen ground extends to a depth of about 427 m (1400 ft). Only a few inches of surface ground thaws in the summer. Exploration Activity Zinc-lead mineralization was

Jan 10, 1984

By H. E. Epstein, B. K. Parekh, W. M. Goldberger

Comminution is the single most expensive operation in mineral processing. It consumes about 50% of the energy required for mineral extraction (Agar, 1976). Current comminution technology is both energy-intensive and inefficient. A novel noncontact comminution process concept was developed in this study, whereby selective lib¬eration of minerals from an ore could be potentially achieved. The process involves application of electric and ultrasonic energy to liberate minerals from gangue particles. Introduction The mineral industry is a large user of energy in the US. Energy usage is increasing as the grade of ores processed decreases. It is es¬timated that comminution of ores uses about 32 000 kWh (115 trillion kJ), or 2% of the electric power produced in the US (NMAB-365, 1981). Among other minerals, ce¬ment, iron, and copper processing plants are the largest users of en¬ergy in comminution. Very little of this energy used in conventional grinding (about 1%) is used to gen¬erate new surface. The remainder is wasted (Table 1). Thus, there are substantial energy-saving and economic incentives to improve the efficiency of crushing and grinding techniques for mineral recovery. With an energy efficiency of only 1%, it would seem possible to devise methods to significantly improve comminution technology. This requires that breakage force be applied only where needed, not indiscriminately as in a conven¬tional ball mill (NMAB-365, 1981). Can apparatus and methods be de¬ veloped for large-scale commer¬cial use that allow energy to be fo¬cused at intergranular bounda¬ries? If this can be done, mono¬mineral grains would remain in¬tact and the grinding action would be selective and substantially more efficient. The novel comminution process concept described in this paper uses a combination of electric and ultrasonic energy. This energy breaks the ore and selectively lib¬erates minerals. The process has also been termed a two-stage or electroacoustical comminution process (Goldberger, Epstein, and Parekh, 1982). The mineral grain boundary is usually the weakest area in an ore. By applying electri¬cal energy to the ore, the rock frac¬tures mostly at grain boundaries. At the same time, the electric shock creates secondary hairline fractures in the ore. Additional application of ultrasonic energy to this ore provides further break¬age. The concept was proven on a molybdenum porphyry ore but needs additional study. Technical Background There have been substantial im¬provements in milling machinery and grinding operations, but rela¬tively few attempts have been made to develop alternatives to conventional impact milling. Sev¬eral methods, however, have been investigated that relate to this study. In the early 1970s there was considerable interest in a size re¬duction method known as the Snyder Process (Cavanaugh, 1972). It involved charging coarse ore into a pressure chamber, pressur¬izing with a gas, and activating a quick-opening (15 msec ) dis¬charge valve that connected the chamber to a discharge duct. This allowed solids to fluidize and ac¬celerate, subjecting the material to a variety of impulse phenomena that caused the desired size reduction. Other nonimpact means of achieving breakage and selective size reduction have been de¬scribed in technical and patent literature. Kanellopoulos and Ball (1975) considered using heat to induce thermal stress in quartz¬ite. This would cause cracking and, thus, increase grinding efficiency. The use of electrical energy to induce thermal stress was studied by the General Elec¬tric Co., in cooperation with the Montana School of Mines, under a research grant from the Anaconda

Jan 9, 1984